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Symposium Proceedings

THE 3RD INTERNATIONAL SYMPOSIUM ON DYNAMIC

HAZARDS IN UNDERGROUND COAL MINES

18-20th July 2018

University of Wollongong, Australia

PRINCIPAL EDITORS

Ting Ren, Linming Dou and Xueqiu He

ORGANISED BY

Centre for Infrastructure Protection

& Mining Safety

University of Wollongong

State Key Laboratory of Coal Resources

& Safe Mining

China University of Mining and Technology

The 3rd International Symposium on Dynamic Hazards in Underground Coal Mines University of Wollongong, Australia

ii

ONLINE PROCEEDINGS OF THE 3RD

INTERNATIONAL SYMPOSIUM ON

DYNAMIC HAZARDS IN UNDERGROUND COAL MINES

18-20th

JULY 2018

UNIVERSITY OF WOLLONGONG

NSW.AUSTRALIA

Principal Editors: Ting Ren University of Wollongong Linming Dou China University of Mining and Technology, China Xueqiu He University of Science and Technology Beijing, China

Organisers:

Centre for Infrastructure Protection & Mining Safety Faculty of Engineering and Information Sciences, University of Wollongong (UOW), Australia Key Laboratory of Deep Coal Resource Mining (CUMT), Ministry of Education China University of Mining and Technology (CUMT), China

Co-organisers: University of New South Wales (UNSW), Australia University of Science and Technology Beijing (USTB), China Anhui University of Science and Technology (AUST), China Chongqing University (CQU), China Monash University (MONASH), Australia Zhong-an Academy of Safety Engineering (ZASE), China


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Organising Committee: Honary Chair

Professor Liang Yuan, AUST

Chair

Associate Professor Ting Ren, UOW

Co-chairs

Professor Linming Dou, CUMT

Professor Ismet Canbulat, UNSW

Professor Xueqiu He, USTB

Jian Zhao, MONASH

Members

Professor Yiyu Lu, CQU

Dr Dennis Black, CoalGas

Professor Yuanping Cheng, CUMT

Professor Sheng Xue, AUST

Professor Jianping Wei, HPU

Associate Professor Alex Remennikov, UOW

Dr Zhejun Pan, CSIRO

Dr Jan Nemcik, UOW

Ms Justine Calleja, UOW

Dr Jianming Wu, UOW

Dr Qianbing Zhang, MONASH

Dr Chen Cao, UOW

Dr Runsheng Lu, UOW

Dr Peijian Jin, UOW

Dr Cunwen Wang, UOW

Dr Cui Ding, UOW

Symposium Secretary

Mr Xiaohan Yang, UOW

Dr Xuwei Li, CUMT

Mr Qingyi Tu, UOW

Working Team

Mr Anxiu Liu, UOW

Mr Jia Lin, UOW

Mr Hongchao Zhao, UOW

Mr Lihai Tan, UOW


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Sponsors:

The organising committee would like to gratefully acknowledge the financial sponsorships of the following organisations:

Supporters:

The organising committee would like to acknowledge the support of the following organisations for this event:

Coal Services The National Facility for Physical Blast Simulation (NFPBS)

NSW RESEARCH ATTRACTION AND ACCELERATION PROGRAM


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Table of Contents

Organising Committee: ............................................................................................................................ iii

Sponsors: .....................................................................................................................................................iv

Supporters: ..................................................................................................................................................iv

Full Papers

Study on failure mechanism and control of extra-large cross section rock burst mining roadway

Yongzheng Wu, Jie He, Yang Wang................................................................................................................ 1

Numerical investigation of the roof weighting and the energy release in a coal mine with super-thick coal

seam and multi-layered hard roof

Binwei Xia, Lang Liu, Yiyu Lu, Lei Zhou ....................................................................................................... 8

Coal and gas outburst and its relationship with tectonic coal

Qingyi Tu, Yuanping Cheng, Ting Ren ......................................................................................................... 14

Comprehensive prevention and control technology of mine pressure bumping in fushun mining area

Guojun Li, Jiquan Sheng ................................................................................................................................ 26

抚顺矿区冲击地压综合防治技术

李国君盛继权 ............................................................................................................................................... 35

Abstracts

Effects of gas pressure on the dynamic compressive strength of coal

Z. Q. Yin, Z. Y. Chen, J. C. Chang, Z. X. HU, H. F. Ma, G. M. Zhao .......................................................... 42

Optimal coal discharge of hydraulic cutting inside coal seams for stimulating gas production: a case study

in pingmei coalfield

Baiquan Lin .................................................................................................................................................... 43

Spatiotemporal multifractal characteristics of electromagnetic radiation in response to deep coal rock bursts

Enyuan Wang ................................................................................................................................................. 44

Numerical simulation study of hard-thick roof inducing rock burst in coal mine

Jiang He, Linming Dou, Songwei Wang, Changhao Shan ............................................................................. 45

Static and dynamic loads superposition mechanism of coal-gas dynamic disaster and its prevention

principles

Linming Dou, Xueqiu He, Ting Ren, Jiang He, Zengyi Wang ...................................................................... 46

Strong-Soft-Strong structural model for roadway preventing rock burst and its realizing technology

Mingshi Gao ................................................................................................................................................... 47


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Residual ground stress affected by the mining order and mining direction: a case study

Wei Yang ....................................................................................................................................................... 48

Regional clustering analysis of coal mine seismic events based on ant colony optimization algorithm

Xiaolin Li, Munezero Eugene, Zhiru Cai ...................................................................................................... 49

Load distribution characteristics and anchoring force fluctuation mechanism of anchoring section under

mining stress

Xigui Zheng ................................................................................................................................................... 50

Dynamic Early-warning of Rock Burst Based on Monitoring of Low-frequency Electromagnetic Radiation

Zhentang Liu ................................................................................................................................................. 51

Research on the dual-porosity/dual-permeability solid-gas coupling model of moisture-containing fractured

coal bed and its flow law

Yuan Zhao, Yong Li, Shugang Cao, Bo Zhang, Hui Wang .......................................................................... 52

Study on precursors of rock instability: insight from acoustic emission (AE)

Jing Li, Qianbing Zhang, Jian Zhao .............................................................................................................. 53


1

Study on failure mechanism and control of extra-large cross section rock burst mining roadway Yongzheng Wu

1, 2, 3, Jie He

1, 2, 3, Yang Wang

1, 2, 3

1. Coal Mining and Design Department, Tiandi Science and Technology Co., Ltd., Beijing 100013, China

2. Coal Mining Branch, China Coal Research Institute, Beijing 100013, China

3. State Key Laboratory of Coal Mining and Clean Utilization, Beijing 100013, China

ABSTRACT

Taking rock burst roadway with extra-large section of Gengcun coal mine in Yima mining district as project

background, geological situation and in-situ stress characteristics of surrounding rock were presented, the in-situ stress

level was high, the strength of coal seam was low and burst potential was high also, extremely thick conglomerate

strata existed in coal seam roof, of which thickness was about 380m, and rock burst appeared intensely. And then

surrounding rock convergence and failure features of rock burst roadway were analyzed, the main influence factors of

rock burst roadway convergence and failure were high in-situ stress, strength dynamic load, pressure relief and

supporting system. Impact damage effect would generated as frequently impact energy, the impact that to surrounding

rock structure and supporting system of rock burst roadway was enormous under some pressure relief, then roadway

would deformation heavily as supporting system failure. The underground industrial test was proceeded in Gengcun

coal mine, the driving section of 13230 upper roadway exceeded 30 m2,so it should be classified as super large section

rock burst roadway, and the following ideas were put forward, which mine pressure characters of rock burst roadway

with extra-large section was obviously different from ordinary roadway that largely of roadway surface convergence

and roof separation were in different position, roadway deformation was faster and more intensely under driving

influence, stable time was longer. And bolts and cables force fluctuated as serration and step shape.

Keywords:

Rock burst; Extra-large section; Mining roadway; Rock bolting; Failure mechanism; Underground testing

Introduction 1.

Rock burst is a common dynamical failure

phenomenon in coal mine underground engineering, and

also surrounding rock deformation and control for rock

burst roadway is a difficulty technology problem in

mining engineer and rock mechanics field.

Many scholars have researched on rock burst

roadway supporting. Kanghongpu et al [1] put forward

supporting selection principle of rock burst roadway,

and then detailed rock bolt supporting designing method

of rock burst was illustrated, a supporting method

combined rock bolt and cable supporting with metal

shed was proposed.

Panyishan et al [2-4] put forward ideas that

supporting strength should be improved and energy

must be absorbed quickly. The following three

conditions that induced to gob-side entry rock burst

were concluded by Jiangfuxing et al [5], which are

about stress, stress gradient and coal mass rock burst

tendency, and then engineering criteria of rock burst

hazard was put forward. Jiangyaodong [6] analyzed

dynamic failure mechanism of coal mine roadway that

induced by blasting.

The failure mode of underground chamber under

blasting condition and reinforcement under different

bolt supporting parameters and blasting resistant results

were studied by Gujincai etal [7]. Strengthening-

weakening-strengthening structural model of rock burst

roadway was put forward by Gaomingshi et al [8-10],

and then supporting parameters was verified by energy

balance theory. Wenjun Ju[11] obtained energy

absorbing index of the main components of rock bolt

supporting, and some field examples were presented.

Dynamic load and deformation characters of

surrounding rock and anchoring body under impact

loading were studied by Hehu et al [12], it pointed that

rock burst supporting body should had initiative yield

function. These research results reference for analysis of

surrounding rock deformation and failure mechanism

and control methods.

Yima coal mine area is a typical rock burst area in

China, some studies and experiments of rock burst

roadway supporting had processed in the area [1, 13-15].

In the paper, to rock burst mining roadway with extra-

large section in Gencun coal mine of Yima coal mining

area, and then geological situation and in-situ stress

were analyzed and mining roadway surrounding rock

deformation and failure characters were analyzed also,

then rock bolt supporting system demand of rock burst

mining roadway were put forward, and also supporting

technology of rock burst mining roadway with extra-

large section was explored.

Geological situation and in situ stress testing 2.

analysis

Geological situation 2.1


2

Gengcun coal mine is a rock burst coal mine, the

testing site is 13230 upper roadway, the average mining

depth is 622m, which located in the east of belt roadway

of the eastern 3rd mining area. It is bordered by mine

property line to the east, mine goaf of 13210# working

face to the north, and coal mass on the south. The 13230

upper roadway is gob-side driving entry with small size

coal pillar, the coal pillar size that to 13210 gob area is

about 6m-8m, and bottom coal that 2m thickness is

layout during driving, the length of 13230 upper

roadway is 1088m, its general arrangement is shown in

Fig.1.

Fig.1. Entry layout of face 13230 in Gengcun Coal Mine

The 2-3 seam is the main mining seam of 13230

working face, its an average thickness is about 10.2m,

and inclination angle is 11°-13°, fractures developed

fully in complicated texture coal seam, about 3-5 layers

mud stone and sandy mud stone dirt bands developed,

its average thickness is about 0.8m, false roof does not

developed within 2-3coal seam, and the roof strata

layers are mud stone and sand stone, and floor is

carbonaceous mud stone and sandstone. Extremely thick

conglomerate strata exist in coal seam roof, its thickness

is about 380m, and distance to coal seam is about 240m,

locating in rock burst hazard area. The conditions of

surrounding rock of roof and floor are shown in Table.1.

Table 1

Conditions of surrounding rock of 2-3 coal top and floor

The burst potential of the 2-3 coal seam was tested

by “Coal Burst Potential Category and Index Measure

Method”, the results for elastic energy index is 16.6,

burst energy index is 9.2, dynamic broken time is 138ms,

uniaxial compressive strength is 20.05MPa, the coal

seam burst potential is strengthening and rock burst

hazard level of 13230 working face is medium.

Surrounding rock in-situ stress testing 2.2

The result of in-situ stress testing with hydraulic

fracturing technique near the 13230 working face, which

included 3 measurement points, is shown in Table. 2.

The first and the second measurement points were

located in the bottom roadway of 13230 working face,

60m from the entrance. σH, σh, σv are maximum

principal stress, minimum horizontal principal stress and

vertical stress respectively.

Table2

In-situ stress measurement results in Gengcun Coal

Mine

NO. Depth

/m /MPa /MPa

V /

MPa

The maximum

horizontal

principal stress

orientation

1 645 13.83 7.29 15.55 N36°E

2 644 12.58 7.09 15.53 N43°E

3 621 14.84 7.69 14.98 N11°E

From Table 2, the following conclusions could be

obtained. Vertical stress in all three measurement points

are greater than the maximum horizontal principal stress,

three is little difference among all of the maximum

horizontal principal stress, mainly dominated by self

gravity.It should be classified as medium stressed area,

stress field is σv>σH>σh, all of the maximum horizontal

stress orientation are NNE.

Some obviously fractures, separation layers and

broken zones developed all in shallow and depth by

investment with borehole observation, some larger

separation appeared within anchoring area in shallow,

which scope is about 2m, the main broken zone in depth

appeared within 5m range in roof.

The coal and rock mass strength testing results

showed as 0-5.1m in roof was 2-3 coal seam, and the

average strength of roof coal body was 15.25MPa, 5.1-

10m was mud stone, and the average strength was

24.44MPa. The upper roadway located near close to the

gob area, and coal body broken after working face

mining, it’s strength variation was larger, the lower

roadway located in coal body, and strength variation

was small, and it’s integrity was obviously better than

upper roadway, the coal average strength of upper

roadway was 14.85MPa.

Surrounding rock deformation and failure of 3.

rock burst mining roadway with large section

The essence reason of rock burst roadway

surrounding rock convergence and failure was

superposition by impact stress wave and surrounding

H h


3

rock stress field, and exceeded the ultimate bearing

capacity of surrounding rock and supporting system, and

then surrounding rock and supporting system was

destroyed [1], it’s main influence were high stress,

strengthening dynamic load ， pressure relief and

supporting system [13]. And mining stress and micro

seismic events were the main phenomenons of high

stress and strengthening dynamic load, pressure relief

means that taking some relief measures had effect on

surrounding rock structure and supporting system, and

supporting system means whether roadway surrounding

rock fiercely convergence could be controlled

effectively and resisting absorbing impact energy.

Micro seismic events 3.1

The burden depth of 13230 working face exceeded

to 600m, in-situ stress level was high, about 380m

extremely thick conglomerate strata existed in coal seam

roof, a large number of elastic energy could be easily

disturbed and then released in a moment, 13230 upper

roadway was gob-side entry with small coal pillar, coal

pillar size was about 6-8m, coal body strength was

lower at coal pillar side, and then failure would easily

appeared after some micro seismic events. Multiple

micro seismic events that energy more than 105J

appeared during working face driving, mining pressure

fiercely and deformation increased rapidly, after micro

seismic events, some part of coal body was broken to

pieces, some local region swell suddenly, and meshes

was teared.

Pressure relief measures 3.2

The combined pressure relief measures that drilling,

blasting and coal seam water injection were applied

during 13230 upper roadway tunneling period.

Two blasting pressure relief boreholes were layout in

heading face, its depth and diameter were 20m and

125mm respectively, when pressure relief scope reached

about20m and then driving distance could be reached

10m in turn, one pressure relief borehole that diameter

125mm was drilling at lower side a separate meter, its

depth was more than 20m, also one water injection

borehole was lay out at lower side of coal wall, layout

type was upper-lower and shallow-depth alternate, and

diameter and depth were 75mm and 80m, respectively,

at last a group three floor cutting blasting boreholes

were lay out every 20m along roadway strike and

vertical to floor.

Pressure relief measures had some huge influence to

surrounding rock structure and supporting system of

rock burst roadway. Firstly, the integrity of certainly

depth scope of coal-rock mass were destroyed by

drilling and blasting, and supporting system was

destroyed at the same time, and also had larger influence

to integrity of rock bolt supporting system. Secondly,

coal-rock mass stress redistribution after pressure relief,

the stress that close to pressure relief region were

released, and stress would redistribute again. Thirdly,

anchoring property of bolts and cables were decreased

after water injection pressure relief, so anchoring force

decreased sharply.

Original supporting system and roadway 3.3

deformation characters

Roadway deformation was seriously for rock burst

roadway in general, larger pre-deformation amount was

set during roadway driving process, most of mining

roadway sections were more than 30m2 in Gengcun coal

mine, strata behavior was violently, the recombination

supporting method that bolt-mesh-cable and 36U shed

and single hydraulic prop supporting was utilized during

roadway tunneling, also hydraulic shed or rock burst

resistant support was applied during mining period, and

bolt interval was 0.7m, roof bolt type was left-handed

rotation screw thread steel reinforcing bar

without longitudinal inner rib , but sidewall bolt was

right-handed rotation fully-thread bolt, and the cable

diameter was 17.8mm, length was 8m, and 36U shed

with interval 700mm was applied.

Under varies supporting pattern, roadway pressure

was fiercely and the largest convergence between two

sides of partly roadway exceeded 3.5m, floor heaving

exceeded more than 2m, some of 36u sheds were

destroyed. and supporting difficulty increased with rock

burst and predeformation , in general roadway section

exceeded 3 m2 was also an important influence factor.

As we know, supporting results decreased with the

disadvantages of right-handed rotation fully-thread bolt,

in additional active supporting effect were

underachieved as low prestress force of bolt and cable,

the primary supporting system would be a failure,

roadway maintain depended on the second level

supporting system, which was about shed, hydraulic

shed or roadway support and so on, so fiercely

convergence of roadway could not be controlled easily.

Rock bolt supporting system of rock burst 4.

mining roadway

The main influence factors of rock burst mining

roadway deformation and failure were high in-situ stress,

strengthening dynamic load, pressure relief and

supporting system, so some corresponding measures

should be adopted, in aspect of high in-situ stress,

surrounding rock stress was decreased or transferred by

some pressured relief measures, in terms of

strengthening dynamic load, some measures were

adopted to absorbing or resistant influence of dynamic

load. While in pressure relief, it took some rational and

effectively relief measures. This paper mainly focuses

on bolt supporting system of rock burst mining roadway.

Bolt supporting is the main supporting type of coal

mine roadway in China [16], which should be adopted

priority in rock burst mining roadway with higher

demands:

(1) Rock burst mining roadway would be influenced

by tunneling and mining as normal roadway, and impact

loading. Hence bolt supporting system should bearing


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not only static loading, but also strengthening loading.

(2) Bolt supporting system should keep integrity of

coal and rock mass. Coal and rock mass and bolt

supporting system should bear impact loading together,

once the coal and rock mass are incomplete, the bolt

supporting system effect would be influenced seriously.

(3) The function of pre-stress have reached

consensus on bolt supporting system [17], an effectively

diffuse of pre-stress should base on high pre-stress and

strengthening surface protection components. Rock

burst roadway would experience intense impact loading

frequently, steel belt, bolt plate and net should keep

enough strength and stiffness with larger size to prevent

surface protection components failure or falling into

coal.

(4) Full length pre-stress anchoring supporting

system should be adopted, avoiding the failure between

the interface of anchoring and non-anchoring of bolt and

cable with impact loading.

(5) Reducing influence of pressure relief to roadway

supporting system. Pressure relief measures should be

adopted out of the range of bolt supporting, drilling

should locate at further site to bolt, water injection

should proceed after bolt supporting.

(6) Bolt supporting system should kept larger burst

resistance, bolt supporting material and its components

should match each other, avoiding failure after larger

impact loading.

(7) Decreasing once effectively supporting and

avoiding roadway repairing. Rock burst accident would

appear easily during roadway repairing, especially

repairing at many regions at the same time in one

roadway. Bolt supporting system should realize once

supporting as far as possible, decreasing and avoiding

roadway repairing.

Field experiments 5.

A field experiment was carried out in 13230 upper

roadway of Gengcun coal mine, which located at rock

burst hazard region. Roadway cross section shape was

three centered arches, tunneling width and height were

7.5m and 4.6m, respectively, classified as extra larger

section rock burst mining roadway. After analysis above,

and combined engineering practical, high pre-stress

strengthening bolt and cable supporting system should

be adopted in 13230 upper roadway, and matching up

the original measures pressure relief. Water injection

must be adopted after bolt supporting.

Roadway supporting scheme 5.1

The rock bolts rolled steel was BHRB500 and its

diameter and length were 22 mm and 2.4m, respectively,

impact absorbing energy more than 40J, and full-length

pre-stress anchoring method was adopted. Surface

protection component W-type steel plate was used for

roof and two sides supporting, its thickness, width and

length were 5mm, 280mm and 450mm, respectively.

Double layers reticules were used for roof supporting,

which iron wire diameter was 3.4mm, bolt supporting

interval and spacing were 0.9m and 0.95m, respectively,

in total fourteen bolts were layout in one row, bolt pre-

stress moment was 400Nm. The cable diameter was

18.9mm, cable in roof and two sidewalls had different

length, about 6.3m and 4.3m, respectively. In total seven

cables were lay out every two rows, the interval and

spacing were 1.9m and 1.8m, respectively, the designing

pre-stress of cable was 200kN. The 13230 upper

roadway supporting layout is shown in Fig.2.

Fig.2. Roadway support layout for upper roadway

13230 in Gengcun Coal Mine

Pressure monitoring 5.2

During roadway tunneling period, three groups of

testing station were lay out in total, and then roadway

surface convergence, roof separation and bolt and cable

force (shown in Fig.4) were monitored.

(1) Roadway surface displacement

Crossing point lay out was adopted to surface

displacement monitoring in three groups of observation

station. In the first station, the maximum displacement

of upper side, lower side, roof and floor were 89mm,

42mm, 192mm, 175mm, respectively, and roof

subsidence and floor heave were the main deformation.

In the second station, the maximum displacement as the

first were 204mm, 103mm, 76mm, 350mm,

respectively, while the maximum displacement in the

third station were 221mm, 180mm, 96mm, 338mm,

respectively. The main deformation appeared at two

sidewalls and floor in the last two stations. From above

monitoring data, surface convergence of the rock burst

mining roadway with extra-large section presented the

following characters: (1) surface convergence data

variation presented more difference in different position

of roadway with distance variation to rock burst source.

(2) Larger floor heave. In view of the service period,

bolt and cable were not installed, so the maximum floor

heave amount reached 350mm. (3) 13203 working face

had not reached to stabilize after mining one year, the

displacement amount at rise side along small coal pillar

obviously was larger than dip side of coal mass. (4)

roadway supporting proceeded under 5.6m thickness top

coal, larger displacement appeared at the roof. (5)

roadway deformation had reached stabilization as

distance to heading face exceed 25m, but floor heave

duration time was long, and it reached stabilization as

distance to heading face exceed 200m. (6) fiercely


5

tunneling influence, roadway section exceeded 30m2,

which would induce fiercely influence, the floor heave

amount exceeded 160mm as tunneling distance reached

to 30m, in total surface convergence amount obviously

was more than that in normal roadway.

(2)Roof separation

The monitoring data of roof separation revealed that

separation value in deep and shallow of the first station

were 135mm and 100mm, respectively. The second

station were115mm and 12mm, and the third station

were 35mm and 19mm. The data presented the

following characters: (1) roof separation data variation

presented more difference in different position of

roadway with distance variation to rock burst source. (2)

roof separation appeared in deep, but the roof separation

value of normal was small and appeared in shallow. (3)

influence fiercely under tunneling, roof separation value

rapidly reached more than 100mm as tunneling distance

reached 30m, it was obviously higher than that in

normal roadway.

(3) Bolt and cable force

The bolt force curves of roof and sidewall is shown

in Fig.3 and Fig.4. The results showed that the pre-stress

of bolt was 40-60KN， individual pre-stress reached

92KN. With driving face advanced, bolt force increased

rapidly in short time, the maximum force reached to

140KN as tunneling distance to 14m. When the distance

that observation station to driving face exceeded 30m,

the entirety blot force stabilized, but a few of that

increased slowly. The maximum bolt force had reached

180KN, which is 10# bolt that located at roadway arch

corner, and approached to yield state.

Another phenomenon appeared during monitoring,

some bolt force appeared variation as zigzag type or step

type. The previous studies [1]

showed that bolt and cable

force of rock burst roadway was different from normal

roadway, bolt force would fluctuate with micro seismic

events with more energy. Although continuous

monitoring was not adopted in this time, daily

monitoring had presented similar principle. Bolt force

all appeared larger fluctuated, as shown in Fig.3 (bolt 7#)

and in Fig.4 (bolt 13)

Fig.3. Roof bolt force curves

Fig.4. Sidewall bolt force curves

Cable pre-stress loss experiment and force

monitoring were proceeded during field testing, cable

pre-stress loss in underground had almost reached to

50%. Sidewall and roof cable force curves are shown in

Fig.5 and Fig.6, respectively, the following conclusions

could be obtained: (1) field cable pre-stress was small

and focused on 70-130KN, it presented more difference

to designing value 200KN，roof cable pre-stress loss

was obviously lower than sidewall one. (2) roof cable

force was larger than sidewall one, an average force of

roof cable could reach to 280KN, but sidewall cable was

only 144KN. (3) Partly cables force fluctuated as zigzag

type or step type under larger micro seismic event, as

bolt 2# and 7# of Fig.5 and 4# shown in Fig.6, but shock

range of roof cable was less than sidewall one.

Fig.5. Sidewall cable force curves

Fig.6. Roof cable force curves

The following obviously differences between rock

burst mining roadway with extra-large section and

normal roadway were presented in the results of

pressure monitoring: (1) roadway surface convergence

and roof separation presented more difference in


6

different region and position under larger micro seismic.

(2) roadway deformation was rapidly and fiercely under

influence of micro seismic and tunneling, roadway

surface displacement and roof separation were several

even decuple times larger than that in normal roadway.

(3) time duration was longer and more stable and for

rock burst mining roadway with extra-large section, it

reached fully stable as roadway driving distance to

200m. (4) bolt and cable force presented as zigzag type

or step type under some larger micro seismic events.

Surrounding rock deformation was controlled

effectively with the supporting scheme. The pressure

monitoring data showed that the maximum roof

subsidence reached to 260mm within range about 100m

of advanced supporting area during mining period, two

sides convergence reached to 510mm, floor heave

reached to 690mm. Roadway floor heave reached to

300-400mm in other regions. Roadway could meet

production demand basically without any maintenance.

As single energy event reached to 107J, bolt and cable

that not broken and roadway kept stabilized was

invested. It showed that bolt supporting system total

presented larger burst resistant, and its fierce

deformation could be controlled effectively.

Conclusions 6.

(1) In-situ stress of rock burst mining roadway of

Gengcun coal mine in Yima coal mining area was high

and self-gravity stress was dominated, coal seam

strength was low and impact tendency was

strengthening, extremely thick conglomerate strata

existed in coal seam roof, which thickness is about

380m, rock burst appeared fiercely.

(2) The main influence factors of rock burst roadway

deformation and failure were high in-situ stress,

strengthening loading, pressure relief and supporting

system. Surrounding rock texture of rock burst roadway

and supporting system would be influenced largely by

pressure relief.

(3) Bolt supporting system requires rock burst

roadway were put forward as following: sufficient burst

resistant, bearing both static loading and strengthening

dynamic loading, coal and rock mass should be kept

integrity, and full length pre-stress anchoring supporting

should be adopted in field, high pre-stress must be

applied and then pre-stress could be diffused effectively,

decreasing influence that pressure relief to roadway

supporting system.

(4) Combined supporting system that high pre-stress

and fully length anchoring strengthening bolt and cable

were adopted with partly pressure relief measures, so

fierce deformation and broken of rock burst roadway

with large section could be controlled effectively.

(5) Pressure features of rock burst mining roadway

with extra-large section were obviously different from

normal roadway. Roadway surface displacement and

roof separation presented larger variation in different

regions and sites, roadway deformation was more

rapidly and more fiercely and stabilization time was

longer under the influence of tunneling, bolt and cable

force fluctuated fiercely as zigzag type or step type.

Acknowledgements

The paper was supported by the State Key Research

Development Program of China (Grant No.

2017YFC0804205).

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Distribution，type， mechanism and prevention of rock-

burst in China[J]. Chinese Journal of Rock Mechanics

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Analysis of blasting tremor impact of roadway stability in

coal mining[J]. Chinese Journal of Rock Mechanics and

Engineering，2005，24 (17)：3131-3136.

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explosion load[J]. Chinese Journal of Rock Mechanics

and Engineering，2008，27(7) : 1315-1320.

[8] Gao Mingshi ， Dou Linming ， Zhang Nong ， et al.

Strong-soft-strong mechanical model for controlling

roadway surrounding rock subjected to rock burst and its

application[J]. Rock and Soil Mechanics，2008，29(2)：

359-364.

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mechanism of roadway supported by bolt-mesh subjected

to rock burst and degree calculation[J]. Journal of Mining

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on roadway support parameters subjected to rock burst

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7

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[18]


8

Numerical investigation of the roof weighting and the energy release in a coal mine with super-thick coal seam and multi-layered hard roof Binwei Xia, Lang Liu, Yiyu Lu

*, Lei Zhou

State key laboratory of coal mine disaster dynamics and control, Chongqing University, Chongqing, China

ABSTRACT

Tongxin is one of the largest coal mines in Datong area in China. As the coal seam is super thick, varying from 14 m

to 20 m, a high stope is created through the top coal caving technology. In addition, the roof is hard with thick layered

sandstone. As a result, great and violent roof weighting appeared frequently. In this paper, a new numerical approach

is used for simulation of longwall mining and investigation of roof weighting. The simulation results of the

progressive mining, the roof damage and caving, as well as the automatic contact of the roof and the cave-in zone are

representative and logical. According to the results, it is found that the energy release can be used to determine which

layers at a given face advance contribute more to the hydraulic support pressure.

Key words:

Longwall mining; Numerical modeling; Roof weighting; Energy release

* Corresponding author. Tel: +86 023-65102421

E-mail address: [email protected]

Introduction 1.

Super-thick coal seams are widespread in Shanxi

province and Inner Mongolia in China. Tongxin is one

of the largest coal mines in north Shanxi. The thickness

of the coal seams in Tongxin varies from 14 to 20 m.

Using long wall mining and top coal caving technology,

the annual coal production from this mine is over 10

million tons. Owing to the extremely high stope, a wide

range of rock-strata are damaged and caved. The main

roof is thick and strong, resulting in extremely violent

roof weighting. This causes various accidents including

extreme compression or failure of hydraulic supports,

pillar instability, water inrush through generated cracks,

significant surface subsidence and wind blast damage.

To prevent such disasters, hydraulic fracturing

technology is proposed as a protective measure, to

extensively and previously break the integrity of the

main roof in this mine area. Hydraulic fracturing is

planned to be operated through a well drilled from the

surface before coal mining. The key layers and the key

positions, that the well will reach for fracturing, remain

to be determined. Thus proper methods should be used

to analyze the temporal movement of the rock strata and

their relationship with the weighing stress.

Nowadays, computer technology is rapidly

developing. Numerical models have been proposed by

various researchers to overcome the limitations of

analytical and physical analog experimental methods.

Numerical models can be classified into two groups:

those based on the discontinuum theory and are solved

using the distinct element method (DEM) [1-3]; and

those based on the continuum theory and are solved by

the finite element/difference method (FEM/FDM). The

continuum model was first used to analyze the stress

distribution under different advance length before

caving [4, 5]. On this basis, Manteghi [6] developed a

model to estimate the first weighting distance according

to the yield elements in the roof. For simulation of

progressive long-wall mining, Singh [7, 8] proposed a

new numerical approach. In this approach, Mohr

Coulomb is applied with tension cut off and strain-

softening to describe the mechanical behavior of the

rock strata. Shear strain and vertical displacement are

used to predict roof caving. If the shear strain or the

vertical displacement of any element exceed a critical

value, the element is deactivated. In addition, hydraulic

power supports are modeled by structure elements.

According to the calculated load on the power support,

the first and the periodic weighting distance are

estimated. The modeled results (weighting distance and

support load) are comparable with field observations,

despite limited face advance because the support effect

of the caving material is not considered. In fact, the

support effect influences the roof weighting and the

stress distribution [9]; the calculated caving height in

Singh's model seems to be unrealistic as it is much

greater than the stope height. For this reason, Norrozi

[10] and Hosseini [11] proposed approaches that

consider the caving material. They used the same

numerical model as developed by Manteghi [6]. They

define an immediate roof to be caved above the coal

seam. After the roof weighting, the yield elements of the

immediate roof and the excavated coal elements are

replaced by cave-in elements. Shabanimashcool [12]

provided a similar method which is more realistic. In his


9

method, the caving height is pre-calculated through the

dilation factor of the cave-in rocks and the stope height.

If the caving height is achieved, then the cave-in

material is introduced into the model. The precondition

for roof caving is that the maximum roof displacement

is greater than the excavation height and the plastic

strain of the roof elements exceed a critical value.

Previous numerical models make important

contributions towards understanding and predicting roof

weighting and strata-support interaction. However, they

should be further developed to provide more effective

means for roof control, especially considering the

relationship between strata movement and weighting

stress during mining advance. In the present paper, a

new numerical approach based on the commercial

software FLAC3D [13] is used for analysis of roof

weighting and strata movement during longwall mining

in Tongxin mine. An energy criterion is used to

determine the key layers.

Theoretical background 2.

In longwall mining, coal seam excavation creates a

void, disturbing the surrounding rock mass around the

void. This causes caving of the immediate roof, damage

and deformation of the far-field rocks, separation and

shear of the bedding plane, gradual roof subsidence and

consolidation of the cave-in material. For the numerical

model, all of these physical processes must be properly

and logically simulated. Thus, an algorithm is proposed

in the present paper. The flow chart of the algorithm is

shown in Fig. 1. The algorithm programmed using the

bulit-in fish language in FLAC3D, contains an outer

loop and an inner loop. The outer loop controls the

advance length. In each loop, the advance length is

increased by △l. Meanwhile, all of the elements in the

volume defined by the increment length and the

excavation height are deactivated, to represent the void

space. The inner loop is the core unit of the algorithm

and it includes four modules: mechanical calculation,

modeling of the roof caving and contact with the cave-in

zone, calculation of the hydraulic support pressure and

calculation of the energy release. In the inner loop, the

mechanical calculation runs until the unbalanced force

ratio achieves a given tolerance value. During the

mechanical calculation, the other three modules are

executed once. The detail of the governing equation can

be found in [14]. According to elasto-plasticity theory,

the released energy is mainly caused by plastic

deformation and can be expressed through Eq. (1).

dd

ddW

plas

ijij

plas

ij

elas

ijij

ijij

(1)

where W is the increment of the released energy [J];

σ is the stress [MPa]; ε is the stain [-]; Ω is the volume

of the rock domain [m3]; elas and plas indicate elastic

and plastic, respectively.

Fig. 1. Flow chart to model the progressive longwall

mining

Simulation of longwall mining in tongxin coal 3.

mine

Model generation 3.1

In the numerical study, panel 8203 in Tongxin mine

is chosen for investigation. A two dimensional

geometrical model (Fig. 2) is generated. The two

dimensional model is valid for simulating the conditions

at the center of the longwall face and can be used to

simulate a representative progressive face advance due

to shorter run time requirement [7]. An advance length

of 600 m with an excavation height of 16 m is proposed

for simulation.


10

Fig. 2. Geometrical model with rock strata distribution in the numerical simulation

Simulation of longwall mining 3.2

Fig. 3 shows the simulation results of progressive

mining, the roof failure and caving, as well as the

automatic contact of the roof and the cave-in zone. It is

clear that the multiple layered roof gradually subsides

and fails, which is representative and logical. The

immediate roof is the first to sustain damage and cave

(0-150 m); the upper layers remain intact. As the

advance length increases, the layers progressively fail

from the bottom up. This suggests that, in addition to the

excavation height, the advance length is another factor

influencing the height of the damage zone.

Analysis of roof weighting and energy release 3.3

The support pressure was measured after an advance

length of 200 m. In the numerical simulation, the

support pressure is recorded and compared with the

measured results in Fig. 4. The results from the model

and the measurement are similar, and reveal that the

support pressure behavior is significantly violent, with

high peaks of over 45 MPa. The distance between two

peaks can be considered as the roof weighting distance.

As shown in figure 4, the average measured and the

modeled weighting distances are about 11.9 and 12.7 m,

respectively.

When comparing the released energy and the support

pressure (figure 5), it is interesting to note that, after the

support pressure reaches a peak, a large amount of

energy is released, leaving the support free from high

pressure. This phenomenon is logical. First, the roof

weighting causes compression of the support, resulting

in high support pressure. Second, the roof is damaged as

it subsides, until it caves and falls to the floor, releasing

energy stored in the rock. Owing to the rock caving, the

bearing stress in the roof above the support is strongly

reduced, and as a result, the support pressure drops.

Fig. 6 shows the temporal development of the

released energy in some thick layers. Since the high

support pressure is related to the high energy release, it

provides a way to determine the key layers for roof

weighting. The contribution of each layer is obtained

from Fig. 6. High energy release occurs in layers 1-8,

suggesting that these layers could the key layers which

must be handled by protective measure. According to

Fig. 6, it is also possible to quantitatively estimate the

upward movement of the rock strata. The upward

movement of the rock failure during face advance is

presented in Fig. 7. Layer 1 is the immediate roof. It

begins to fail after an advance length of 30 m. After an

advance length of 57 m, layer 2 begins to fail. It is

important to note that the (34 m thick) rock strata

between layers 2 and 5 are also damaged when layer 2

begins to fail, indicating that layer 2 controls the

deformation of these formations. According to the same

principle, layer 5 begin to fail after an advance length of

288 m and controls the deformation of 48 m thick rock

formation above it. Since layers 8-10 have little

contribution to the released energy, it can be concluded

that layers 2 and 5 are the key layers for roof control.

Hydraulic fracturing operations should be used in these

two key layers. As the failure of the two layers begin at

different face advance, the positions of the operation

must be considered separately. Before an advance length

of 288 m face advance, layer 5 is not damaged. This

prevents the weighting of the above rock formation, thus

the integrity should be maintained. Fracturing operation

is unnecessary in this section.


11

Fig. 3. Simulation results of the progressive caving and roof failure at different advance lengths

Fig. 4. Support pressure modeled numerically compared with that measured


12

Fig. 5. Released energy compared with support pressure

Fig. 6. Temporal development of the released energy in some thick layers


13

Fig. 7. Upward movement of the rock strata during face advance

Conclusion 4.

In the present paper, a new numerical approach was

used for simulation of longwall mining. According to

the analysis of the energy release in each layer, it is

found that some thick layers control the deformation and

failure of rock formations above them. These layers can

be considered as key layers which should be handled by

fracturing operation to reduce the weighting stress.

Since the failure of the key layers begins at different

advance lengths, positions of the fracturing operation

can be more effectively planed. According to the

simulation results, layers 2 and 5 are the key layers.

Before an advance length of 288 m face advance, layer 5

is not damaged. This prevents the weighting of the

above rock formation, thus the integrity of layer 5

should be maintained. The fracturing operation should

be conducted at a position more than 288 m to the panel

entry.

Acknowledgments

The work presented in this paper is funded by the

China Postdoctoral Science Foundation (grant number

2016M602655); and the Chongqing postdoctoral

scientific research project (grant number Xm2017199)

References

[1] Wang Fangtian, Tu Shihao, Yuan Yong, Feng Yufeng,

Chen Fang, Tu Hongsheng. Deep-hole pre-split blasting

mechanism and its application for controlled roof caving

in shallow depth seams. Int J Rock Mech Mining Sci.

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Haohao. Caving mechanisms of loose top-coal in

longwall top-coal caving mining method. Int J Rock

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[3] Gao Fuqiang, Stead Doug, Coggan John. Evaluation of

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[4] Qian Minggao. The retrospect of practice and theory for

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[6] Manteghi H, Shahriar K, Torabi R. Numerical modelling

for estimation of first weighting distance in longwall coal

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[7] Singh GSP, Singh UK. A numerical modeling approach

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performance of hydraulic powered support in longwall

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[8] Singh GSP, Singh UK. Numerical modeling study of the

effect of some critical parameters on caving behavior of

strata and support performance in a long wall working.

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H. A model for determining the breaking characteristics

of immediate roof in longwall mines. Yerbilimleri. 2012;

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[11] Hosseini Navid, Goshtasbi Kamran, Oraee Behdeen,

Gholinejad Mehran. Calculation of periodic roof

weighting interval in longwall mining using finite

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42(09):2235-44. (in Chinese)


14

Coal and gas outburst and its relationship with tectonic coal Qingyi Tu

1,2, Yuanping Cheng

1,2*, Ting Ren

3

1. Key Laboratory of Coal Methane and Fire Control, Ministry of Education, China University of Mining and

Technology, Xuzhou 221116, China

2. Faculty of Safety Engineering, China University of Mining and Technology, Xuzhou 221116, China

3. School of Civil, Mining & Environmental Engineering, University of Wollongong, NSW 2522, Australia

ABSTRACT

Coal and gas outburst is a kind of micro-geological disaster that is controlled obviously by tectonic movement.

Combined with the statistics of tectonic types and coal characteristics for outburst cases, the relationship between

outburst and tectonic coal is obtained, after the comparing of the tectonic coal and intact coal and summarizing of the

outburst basic characteristics. Then, this relationship is demonstrated from an energy perspective. Tectonic coal is a

product of tectonic movement, and the primary structure of tectonic coal is destroyed by tectonic process. Due to the

properties of low strength, weak cohesion and fast gas release speed, tectonic coal is prone for outburst. The coal

particle or pulverized, with a size less than 1mm, is the guarantee for rapid desorption of adsorbed gas, which widely

exists in outburst coal. However, the generation of these coal particles produce a large amount of newly added surface

area, require a lot of crushing work. The estimation results show that the crushing of coal expends the vast majority of

outburst energy, when intact coal outburst occurs. The stress required for tectonic coal crushing is 16.51 MPa, while

the stress required for intact coal crushing reaches 736.70 MPa. Meanwhile, under the condition of tectonic coal, the

corresponding seismic magnitude of 5 outburst cases are 2.07 ~ 2.93; however, the corresponding seismic magnitude

are 3.85 ~ 4.39, when intact coal outburst occurs. Obviously, only the outburst energy requirement of tectonic coal can

be satisfied in current mining depth

Key words:

Coal and gas outburst; Tectonic coal; Tectonic movement; Outburst energy; Seismic magnitude

* Corresponding author. Tel: +86 0516-83885948

E-mail address: [email protected]

Introduction 1.

Coal and gas outburst (hereinafter referred to as

outburst) is an unexpected disaster for underground

mine1, 2

, and is an instantaneous releasing process of gas

and stress energy under the combined action of

geological factors and mining activities, which emits a

large amount of coal (rock) and gas into the production

space in a short time. A huge dynamic effect occurs,

destroying the roadway facilities and ventilation system,

and causing worker injury and death3. More than ten

thousand of cubic meters of gas cause the worker

suffocation and air backflow, even induce secondary

disasters such as gas combustion and gas explosion4.

Over the years, many scholars have conducted

extensive studies on outburst process and mechanism

based on on-site investigation, similar experimental

simulation and theoretical analysis. However, multiple

factors affect this outburst hazard, and there are complex

coupling relationships between these factors, so that the

mechanisms of outbursts under various geological and

mining conditions are have not been fully understood5, 6

.

Hanes et al.7 proposed that both stress and gas play an

important role in outburst, but one of them maybe is

dominant factor for an outburst. Paterson8 believed that

outburst is a coal (rock) structural instability process,

and established a mathematical model for outburst. Sato

and Fujii analyzed2 the seismic waves of a large-scale

outburst in Sunagawa coal mine; they noted that the

earthquake could induce outburst and played a major

role in this outburst. Jiang and Yu explained the outburst

process with the view of spherical shell instability. This

hypothesis is that coal failure caused by stress is only a

necessary condition for outburst, and the sufficient

condition is crack expansion and shell instability under

the effect of gas pressure9. Moreover, Hu et al.

10

suggested that outburst is a mechanical failure process,

divided into four stages including preparation, trigger,

development and termination.

The abovementioned studies confirm that outburst is

the comprehensive result of stress, gas and coal

properties11

. These studies mainly focuses on the

contribution of stress and gas to outbursts, but lacks

attention to the influence of coal properties on outburst,

so that these have failed to answer that what kind of coal

can lead to outburst. Generally, outburst preparation is a

geological process that takes a long time, and the coal's

primary structure has destroyed under the effect of

tectonic movement12, 13

. Hodot11

found that the structural

damage coefficient of outburst risk coal seam, at central

zone in Donbass of the Soviet Union, is 3 times that of

the no outburst risk coal seam, and a large amount of


15

tectonic coal was found in outburst risk coal seam.

Shepherd et al.1 confirmed that thrusts, strike-slip faults

and recumbent fold hinges are especially outburst prone.

Such areas are usually distributed with “tectonic coal”.

This work, the properties of tectonic coal are

compared with the intact coal, and the basic

characteristics of outburst are summarized; then,

combined with the statistics of tectonic types and coal

characteristics for outburst cases, the relationship

between outburst and tectonic coal is obtained. Using

the in-site outburst data, the newly added surface area of

outburst coal, coal crushing work and outburst energy

for several typical outburst cases are calculated; and

based on these, the stress demand and seismic

magnitude for outburst are analysis, which will be used

to demonstrate the relationship between outburst and

tectonic coal.

Comparison of tectonic coal and intact coal 2.

Coal is a solid combustible organic rock, which is

transformed from plant debris with long and complex

biochemical, physical, chemical and geochemical

processes14

. In addition to the influence of sedimentary

factors, the formation and distribution of coal seams

have undergone complicated structural changes in the

later period, resulting in the complexity and diversity of

occurrence of coal seams15

. Generally, there are two

existence types of coal body in coal seam. The first type

of coal is not affected or less affected by tectonic

movements, thus maintains its primary structure. This

type of coal is relatively complete with clear surface

fractures and more edges and corners, as shown in Fig

1a and Fig. 2a. However, the another type of coal is

obviously affected by tectonic movements, and the

primary structure of coal is destroyed. So that, the coal

is cut into fine particles, and then coal powder forms.

However, there is a reconstruction effect under the

tectonic stress, which causes the compact combination

of coal powder. Finally, this type of coal forms, but the

coal joint is unclear or loses its meaning, as shown in

Fig 1b and Fig. 2b. In this work, the former is defined as

intact coal, while the another is defined as tectonic coal.

According to the classification of coal damage types in

the “specification tor indentification of coal and gas

outburst mine” (AQ 1024-2066)16

, type I and II coal

belong to intact coal, while type III, IV, and V show the

characteristics of tectonic coal. Moreover, it is worth

noting that the intact coal is only relative.

Fig. 1. Comparison of intact coal and tectonic coal in 21 coal seam, Guhanshan mine

(a) Intact coal; (b) Tectonic coal

Fig. 2. SEM images of tectonic coal and intact coal in 21 coal seam, Guhanshan mine

(a) Intact coal; (b) Tectonic coal

The strong Neopaleozoic modification of multistage tectonization on coal seam further affects pore


16

structure17, 18

. Some scholars17, 19

confirmed that the

specific surface area and total pore volume in larger

pores increased greatly under the condition of intense

tectonism, while specific surface area and total pore

volume in smaller pores changed more complex. These

may suggest that tectonism deeply promotes the damage

of macromolecular chain and aromatic layer in coal

structure, resulting in the disorderly development in

molecule structure of coal, that is, an increase in specific

surface area and total pore volume20

.

Due to the controlling of pore structure on gas

storage and migration, the changes of pore structure will

directly affect the gas storage and migration in coal

seam11

. Based on the test of methane diffusion

coefficient in tectonic coal and intact coal with the same

location, the methane diffusion coefficient of the

tectonic coal formed by tectonism is generally higher

than that of the intact coal21

. In addition, affected by

tectonic action, the fragmentation degree of tectonic

coal increases, which may lead to the increase of gas

migration channels22

. Therefore, it is easier for tectonic

coal to reach adsorption equilibrium state23

. Meanwhile,

the tectonic coal has a larger initial gas desorption

capacity, and the average desorption rate of tectonic coal

is much larger than that of intact coal at the first minute,

which is very important for outburst24

.

In addition, tectonic coal shows a low strength and

weak cohesion properties, it is highly broken even

powdered after the disturbance of mining activities.

Therefore, it is difficult to obtain a relatively complete

tectonic coal from in-site, and it is impossible to achieve

a standard specimen (with a size Φ50×100mm) using

the core method, which is widely used in intact coal25

.

Despite this, the uniaxial compressive strength of

tectonic coal, obtained by the reconstituted coal

specimen test26

or point load test27

, was usually less than

4 MPa26, 28, 29

. While, most of intact coal’s uniaxial

compressive strength was higher than 10 MPa30, 31

.

In summary, these obvious differences between

tectonic coal and intact determine that tectonic coal has

a higher outburst risk than intact coal.

Basic characteristics of outburst 3.

Since the first reported outburst incident occurred in

the Issac coal mine, Loire coal field, France in 1834,

more than tens of thousands outbursts have occurred in

the world. Although the geological backgrounds and

mining conditions vary widely, but there are some basic

characteristics for these outbursts.

Outburst is an instantaneous process, accompanied

by a strong dynamic effect. Generally, outburst only

lasts a few seconds or tens of seconds, e.g. the

Zhongliangshan mine outburst, on November 4th of

1977 in Chongqing, lasted for 39s32

. Table 1

summarizes the outburst coal (rock) mass and

distribution length for 5 typical outburst cases. It can be

found that hundreds to thousands tons of outburst coal

(rock) were thrown out with a distribution length about

tens to hundreds meters. Moreover, there is an outburst

hole in the coal seam after outburst, and this hole

usually shows a characteristic of smaller mouth and

larger cavity. However, due to the difference of outburst

condition, some outburst holes clearly retain in coal

seam (Fig. 3a), while some of other outburst holes are

partially or completely filled by outburst coal (rock), as

shown in Fig. 3b.

Table 1

Coal and gas outburst cases

Outbur

st case Date

D

epth

of

cove

r (m)

Out

burst

coal

mass

(t)

Distrib

ution

length (m)

Daping

coal mine

2004/

10/20

6

12

189

4 256

Xiangs

hui coal

mine

2012/

11/24

2

03 490 66

Macha

ng coal

mine

2013/

03/12

6

70

205

1 188

Bailon

gshan coal

mine

2013/

09/01

5

00 868 200

Fifth

mine of

the

Yangmei

block

2014/

05/13

8

00 325 30

Fig. 3. Outburst hole for outburst cases


17

(a) Bailongshan mine outburst, China; (b) Xinxing mine outburst, China

According to the in-site investigation, there is

obvious zonal distribution for outburst coal (rock).

Interestingly, the rock with a large size usually

distributed near the outburst point, while the coal

concentrated in the area far from the outburst point.

Fig.4 shows the distribution feature of outburst coal

(rock) in Xinxing mine outburst32

, this accident occurred

in the exploring roadway of the No. 15 coal seam, and

the outburst point located in large fault zone. The

outburst coal (rock) mass was 3,845 tons, including

1,697 tons of coal and 2,148 tons of rock. The

distribution length of the coal (rock) in the roadway was

317 m, including a length of 278 m in the exploring

roadway. As shown in Fig. 4, the exploring roadway

was piled up by large rocks inside of Section J-J, and the

largest rock was found with a size of 3.6 × 3.0 × 2.6 m;

then, within the range of 11 m (Section J-J) to 113 m

(Section F-F), it was filled with rock; while the coal

concentrated in the range of 113 m (Section F-F) to 278

m (Section A-A), and a large amount of coal was

powdery.

Fig. 4. Distribution feature of outburst coal (rock) in Xinxing mine outburst

32

In many cases, outburst coal is highly crushed or

even to pulverized, and the accumulation angle for

outburst coal is less than its repose angle, e.g. Fig. 5a

shows the particle size characteristics in Fifth mine of

the Yangmei block outburst; while, Fig. 9b shows the

coal pulverized that existed during the Machang

outburst. Hu et al.33

conducted an exploration on the

coal particle composition in several outbursts, and the

weight distribution of different particle size ranges is

shown in Table 2. It can be concluded that the outburst

coal contained a large fraction of coal particles (d < 10

mm), accounting for more than 60%. For the sizes that

are less than 1 mm, this fraction still reached 30%.

Moreover, the coal pulverized, with a particle size less

than 0.1 mm, was also found in these outbursts,

accounting for 4.6%, 3.6%, 25.4%, 4.3%, 3.5%, 6.6%

and 1.1%, respectively.

Fig.5. Particle size characteristics of outburst coal

(a) Fifth mine of the Yangmei block outburst; (b) Machang mine outburst


18

Table 2

Wight distribution of different particle size ranges of outburst coal33

Outburst location

Diameter (mm)

< 0.1

mm

0.1 ~ 1.0

mm

1 ~ 5

mm

5 ~ 10

mm

>10

mm

# 4 coal seam in Yutianbu mine, +90 m elevation 4.6 39.1 23.5 16.0 16.8

# 4 coal seam in Yutianbu mine, +90 m elevation 3.6 38.0 31.0 17.6 9.8

K2 coal seam at the open-off cut of the 5th cross-cut in

Zhongliangshan mine, +280 m elevation 25.4 26 27.6 1.0 20.0

K1 coal seam at the Xisi half rising cross-cut in

Zhongliangshan mine, +280 m elevation 4.3 29.9 24.6 14.1 27.2

K10 coal seam at the 5th cross-cut in Zhongliangshan mine,

+280 m elevation 3.5 30.4 30.5 19.4 16.2

K10 coal seam at the Xisi half rising cross-cut in

Zhongliangshan mine, +280 m elevation 6.6 27.5 16.9 18.2 30.8

# 6 coal seam in Nantong mine, -100 m elevation 1.1 11.9 23.9 23.5 39.5

Average value 7.0 29.0 25.4 15.7 22.9

Outburst need a large volume of gas to be generated

instantaneously to produce a large transport energy,

transporting the outburst coal into the mining space. In

the investigations of outbursts, the gas emission amount

usually take as gas exhausted during the period

beginning from the moment of concentration rising to

the time when the gas has returned to the normal level.

The wholes process lasts for tens of hours even to

several days, and the gas emission amount reaches tens

of hundreds of thousands cubic meters. As shown in Fig.

6, the gas emission amount of Daping mine outburst and

Machang mine outburst were 249500 m3 and 352000 m

3,

respectively. However, only a small proportion of gas

emission amount is effective and provides gas energy

for outburst in the short time of outburst.

Fig. 6. Variation in gas concentration over time for outbursts

(a) Daping mine outburst; (b) Machang mine outburst

Relationship between tectonic coal and outburst 4.

As listed in Table 3, Gray34

summarized 105

outburst cases from the 8 coal-producing countries.

Among which, 87 outburst cases identified the tectonic

types, and most of these outbursts occurred in the

tectonic zone such as fault, fold and heavily structured

zones. Table 3 shows that about 95.4% of 87 outbursts

occurred in the tectonic zone; more specifically, the

proportion of fault, fold and heavily structured zones are

47.13%, 36.78% and 11.49% respectively. Only 3

outburst cases, occurred in Bowen Basin of Australia,

were found no relationship between tectonic action and

outburst. However, in these three cases, the outburst

coal (rock) mass were less than 30t, and distribution

length were only a few meters. The intensity for these

three outbursts were much smaller than other cases.

Therefore, there is a close relationship between the

tectonic movement and outburst, especially for the

large-scale outburst.

The influence of tectonic movement on outburst not

only reflects in the changes of coal seam stress and gas,

but also in the destroys of coal primary structure, thus

results that a large amount of tectonic coal exist in

outburst zone. Table 4 lists the protodyakonov

coefficient ( f ) of coal in outburst zone, collecting

from several cases in China. The protodyakonov

coefficient ( f ) of coal is generally low with the

maximum value of 0.33 in these cases. Obviously, the

strength of coal in these cases is more conformed to the

tectonic coal.

0 2000 4000 60000

10

20

30

40

Gas

con

centr

atio

n (

%)

t (min)

Machang mine outburst

Totial gas release: 352000 m3

0 200 400 600 8000

5

10

15

20

Daping mine outburst

Totial gas release: 249500 m3

Gas

conce

ntr

atio

n (

%)

t (min)

(a) (b)


19

Table 3

Outburst cases and the tectonic types in the world34

Tectonic type Fault Thickness

changes Fold

Heavily structured

zones

No specific

structures

No

clear

Russia 9

10

Kazakhstan 4

2

1

China 3

20 3

5

Turkey 8 1 5 6

Britain

1

Japan 1

1

New Zealand 5

1

Sydney Basin,

Australia 9

1 1

Bowen Basin,

Australia 2

1

3 2

Table 4

Protodyakonov coefficient ( f ) of coal in outburst cases, in China34

Number Mine Date Protodyakonov coefficient ( f )

1 Daping mine 2004.10.20 0.12

2 Wangfenggang mine 2006.01.05 0.29

3 Malingshan mine 2006.01.20 0.17

4 Jiulishan mine 2011.10.27 0.17~0.19

5 Xiangshui mine 2012.11.24 0.33

6 Bailongshan mine 2013.09.01 0.3

After many years of mining activities, it has been

observed that tectonic coal were widely distributed in

most coalfields in Southwest and most parts of

Northeast in China, such as Guizhou Province, Yunan

Province, Huainan and Huaibei coalfield in Anhui

Province35

, as shown in Fig. 7. Then, the 31 outburst

cases in Table 4, which occurred in China, are marked

in Fig. 6. It can be found that these outbursts mainly

occurred in the tectonic coal distribution zones,

especially in Southwest of China, such as Yunnan,

Guizhou and Sichuan Province. In fact, more outbursts

have been reported in these zones. Therefore, it can be

concluded that there is a direct relationship between

outburst and tectonic coal.

Fig. 7. Distribution of tectonic coal and outbursts in China


20

Demonstration of the relationship between 5.

outburst and tectonic coal

Outburst energy 5.1

Significant researches conducted by Hodot has

provided a detailed analysis of outburst energy, and he

noted that the outburst comes from stress energy, gas

energy and mining activities additional energy11

. As

shown in the following equation, it is generally

concluded that most of outburst energy is dissipated in

the form of work during the outburst process, resulting

in the crushing and transport of outburst coal; and the

remaining energy is transferred to vibrations energy,

sounds energy, etc.

1 2 3e g fW W W W W W (1)

where eW is the stress energy, MJ; gW is the gas

energy, MJ; fW is the mining activities additional

energy, MJ; 1W is the crushing work, MJ; 2W is the

transport work, MJ; and 3W is the remaining energy,

MJ. However, the mining activities additional energy

( fW ) and remaining energy ( 3W ) are ignored in many

studies.

1 2e gW W W W (2)

During the outburst process, the rapid transfer of

stress causes the failure of coal, is accompanied by the

release of stress energy36

. The stress energy of coal,

under the triaxial stresses condition, can be expressed as: 2 2 2

1 2 3 1 2 2 3 3 1

12

2eE

E

(3)

where eE is the stress energy for unit volume,

MJ/m3;

1 , 2 and

3 are the triaxial stresses, MPa;

E is the elastic modulus, MPa; and is the Poisson’s

ratio. Based on the Eq. (3), the smaller elastic modulus

is, but the larger stress energy is. Generally, the elastic

modulus of coal is lower than that of the surrounding

rock about 1 ~ 2 orders of magnitude. Therefore, the

stress energy mainly comes from the coal body.

Gas is one of the mainly power sources for outburst,

and plays an important role in the transport of outburst

coal. Gas contributes outburst by the release of gas

expansion energy. However, gas expansion occurring

over a short period can be simplified to an adiabatic

process.

gdW PdV (4)

where P is the gas pressure, MPa; V is the

gas volume, m3. There is a relationship

between gas pressure ( P ) and gas volume

(V ).

1 1 2 2PV PV (5)

where is the adiabatic coefficien. When the

gas pressure changes from 1P to

2P , the gas

energy is

2

1

1

2 2 1

2

( ) 11

P

g gP

PV PW dW PdV

P

(6) Outburst requires a large amount of gas, including

free gas and adsorbed gas, to provide outburst energy.

Among them, free gas is stored in the fracture system

and exists in gas phase, which can immediately

participate in outburst. 1

1

2

( )f

m

PV V

P (7)

where fV is the free gas volume, m

3; is the coal

porosity, %; mV is the coal volume, m3. However, the

process of adsorbed gas participate in the outburst is

more complex. Desorption of adsorbed gas is the

premise condition for this process, while desorption of

gas from the coal surface is considered as instantaneous

process. Considering that the broken of coal causes great

damage to the gas seepage channel, the gas diffusion in

pore system, which are affected by particle size and

diffusion coefficient, becomes the controlling factor for

the process of adsorbed gas participate in the outburst.

Based on the unipore diffusion model37, 38

, the gas

desorption volume in short time can be calculated by. 1

22

12( )tQ D

tQ d

(8)

where d is the diameter of coal particle, m; D is

the diffusion coefficient, m/s; t is the diffusion time, s;

tQ is the gas desorption volume at time t , m3; and Q

is the final gas desorption volume, which is

(1 )aad ad

a

abPabPQ M A

a bP a bP

(9) where

aP is the atmospheric pressure, MPa.

Therefore, the adsorbed gas volume can be expressed as a

itV Q (10)

where aV is the adsorbed gas volume, m

3;

itQ is

the gas desorption volume in i particle size range, m3.

Based on the Eq. (7) and Eq. (10), the Eq. (6) changes as 1

2 12 2

2

( ) ( ) 11

f a

g

P PW V V

P

(11) As the previous analysis, the crushing of coal has a

great significance for the quick desorption of adsorbed

gas, while is requires a large amount of crushing work to

overcome the surface energy.

1W S (12)

where is the crushing work ratio, J/m2; S is the

newly added surface area, which is

6 6i

im m

G GS

d d

(13)


21

where G is the outburst coal mass, kg; i is the

mass proportion of coal in i particle size range, %;

is the density of coal, which is considered as a constant,

kg/m3; and md is the average diameter of coal before

outburst, m.

Assuming that only horizontal displacements exist,

and then the whole distribution length of outburst coal is

divided into several small segments, so that the transport

work in each segment can be expressed by: 6

2 ( cos sin ) 10i i m iW m g f a a l m (1

4) where

2iW is the transport work of No. i segment,

MJ; im is the outburst coal mass of No. i segment, kg;

il is the distance from outburst point, m; mf is the

friction coefficient; g is the gravity acceleration, m/s2;

and a is the coal seam angle, º. Therefore, the total

transport work is

2 2iW W (15)

Calculation of outburst energy 5.2

The development of outburst requires two conditions,

including crushing and transport of outburst coal.

Among them, the transport of outburst coal provides the

space condition for coal crushing, while the crushing of

outburst coal supplies the gas for coal transport.

Therefore, the outburst coal is highly broken or even to

powdered, and the proportion of small size coal particles

(d< 1 mm) reached 30%, as shown in Table 3. Yang39

obtained a modified relation between particle size and

gas initial desorption rate of coal, as following:

1 2

2 1

v r

v r

(16)

where 1v and 2v are the gas initial desorption rate

of coal particles with the corresponding coal radius 1r

and 2r , respectively; and is the correction factor. It

is easy to conclude that the gas initial desorption rate

decreases with the particle size showing a logarithmic

variation, so that the reduction of coal radius causes the

increasing of initial desorption rate40

. Yang39

and

Busch41

found that there is a critical diameter for the

desorption of coal particles, which is usually millimeter-

scale. When the coal particle is larger than this critical

diameter, the rate of gas desorption slowly decreases

with the particle size. However, when the coal particle is

smaller than this critical diameter, the gas initial

desorption rate increases sharply with their sizes' decline.

Therefore, the existence of small size coal particles (d<

1 mm) in outburst coal is a guarantee for the rapid

desorption of adsorbed gas during the outburst process.

A large amount of outburst energy is needed for the

crushing and transport of coal. Based on the previous

analysis, the crushing work and transport work of the 5

outburst cases in Table 1 will be calculated; then, the

demand of outburst energy will be obtained for these

cases. As the division method in Table 3, the particle

size distribution of outburst coal are also divided into 5

particle size ranges, and the proportion of each range

takes the average value in Table 3. Moreover, the

particle size of coal in each range is characterized by the

average particle size except for the particle size range

(d >10 mm), while the particle size is simplified to

10mm in this range. Meanwhile, the coal body

undergoes a serious destruction during outburst.

Therefore, the influence of the original surface area of

coal on the newly added surface area is limited, which

can be ignored. Using the Eq. (13), the newly added

surface area in each particle size range for 5 outburst

cases is calculated, as listed in Table 5. The crushing or

pulverization process of coal produces a large amount of

newly added surface area. Especially, the generation of

pulverized coals (d < 0.1mm) have the greatest impact

on the newly added surface area, and these pulverized

coals create 68.1% of the newly added surface area.

Although the proportion of coal particles, which the

particle diameter are larger than 10 mm, reach 22.9%,

but these coals have little effect on the newly added

surface area, which only accounts for 1.1% of the total

newly added surface area.

Table 5

Newly added surface area in each particle size range for

5 outburst cases

Outbu

rst

case

Newly added surface area (m2)

< 0.1

mm

0.1 ~

1.0

mm

1 ~ 5

mm

5 ~

10

mm

>10

mm 合计

Dapin

g coal

mine

1136

4000

4279

948.1

6872

51.4

1699

18.9

1858

82.6

16687

001

Xiang

shui

coal

mine

2940

000

1107

272.7

1778

00

4396

0

4809

0

43171

22.7

Macha

ng

coal

mine

1230

6000

4634

727.3

7442

20

1840

04

2012

91

18070

242.3

Bailon

gshan

coal

mine

5208

000

1961

454.5

3149

60

7787

2

8518

8

76474

74.5

Fifth

mine

of the

Yang

mei

block

1950

000

7344

15.6

1179

28.6

2915

7.1

3189

6.4

28633

97.7

As mentioned above, there are two existence types

of coal body in coal seam, including tectonic coal and

intact coal. Based on previous studies, coal type is one

of the most important factor for outburst. Therefore,

there is a great difference between tectonic coal and

intact coal in the outburst conditions, especially in the

demand of outburst energy. In this work, the crushing

work ratio of tectonic coal and intact coal are selected,

respectively; and then, the crushing work for these two

types are calculated when outburst occurs. The crushing


22

work ratio of tectonic coal referenced the experimental

results conducted by Cai et al.42

, and 7 tectonic coal

samples (Protodyakonov coefficient f was 0.119 ~

0.30) were used in this experiment. The results shown

that the crushing work ratio of tectonic coal was 10.7 ~

28.8 J/m2, with an average value of 19.9 J/m

2. While,

Luo et al.43

. tested the crushing work ratio of intact coal

(Protodyakonov coefficient f was 0.918) was 1.25×104

~ 1.45×104 J/m

2, with an average value of 1.32×10

4 J/m

2.

In addition, Eq. (14) is used to calculate the transport

work of each outburst case. To simplify the calculation,

the distribution of outburst coal is regarded as a uniform

distribution, and the friction coefficient mf is 0.532

.

Finally, the crushing work and transport work of these 5

outburst cases are calculated; then, the demand of

outburst energy for each case is obtained, as shown in

Table 6.

Table 6

Calculation results

Outbur

st case

Tectonic coal Intact coal

Crus

hing

work

1W

(MJ)

Tran

sport

work

2W

(MJ)

Outb

urst

ener

gy

W

(MJ)

Crush

ing

work

1W

(MJ)

Tran

sport

work

2W

(MJ)

Outb

urst

energ

y W

(MJ)

Dapin

g coal

mine

332.

07

1212

.16

1544

.23

2202

68.41

1212

.16

2214

80.57

Xiangs

hui

coal

mine

85.9

1

80.8

5

166.

76

5698

6.02

80.8

5

5706

6.87

Macha

ng

coal

mine

359.

60

963.

97

1323

.57

2385

27.20

963.

97

2394

91.17

Bailon

gshan

coal

mine

152.

18 434

586.

18

1009

46.66 434

1013

80.66

Fifth

mine

of the

Yang

mei

block

56.9

8

24.3

6

81.3

4

3779

6.85

24.3

6

3782

1.21

The results show that both tectonic coal outburst and

intact coal outburst expend large amount of outburst

energy for the crushing and transport of coal. The

outburst energy of tectonic coal outburst is mainly used

to provide the transport work, and the transport work is

2 ~4 times of the crushing work. However, the crushing

of coal expends the vast majority of outburst energy,

when intact coal outburst occurs; the crushing work is

181 ~ 1551 times of the transport work. Considering

that the transport of coal is powered by gas, while the

crushing of coal is leaded by stress6, 44

, the difference of

crushing work indicates that crushing work of intact

coal outburst can only be satisfied under extremely high

stress condition.

Based on the assumption that the three principal

stresses are equal (1 2 3= = ), Eq. (3) is used to

estimate the stress, which is required for the crushing of

coal in outburst process. Moreover, some parameters are

needed for the estimation, e.g. the elastic modulus and

Poisson's ratio of tectonic coal are 0.5 GPa and 0.35,

respectively; while, the elastic modulus and Poisson's

ratio of intact coal are 2.0 GPa and 0.30, respectively.

These parameters referenced the recently published

literatures30, 45, 46

, ensuring that the values are reasonable.

Table 7 lists the estimation results, among which, the

stress required for the crushing of tectonic coal is 16.51

MPa, while the stress required for the crushing of intact

coal is much higher than that of tectonic coal, reaching

736.70 MPa. As listed in Table 2, the maximum cover

depth of these 5 cases is 800 m, and the minimum depth

is only 203 m. Obviously, the ground stress in these

cover depths maybe satisfy the requirement for tectonic

coal outburst, but it is almost impossible to reach the

requirement for intact coal outburst.

Table 7

Stress and seismic magnitude for 5 outburst cases

Outburst

case

Tectonic coal Intact coal

1 2 3= =

(MPa)

Magnit

ude 1 2 3= =

(MPa)

Magnit

ude

Daping

coal

mine

16.51 2.93 736.70 4.36

Xiangsh

ui coal

mine

16.51 2.28 736.70 3.97

Machan

g coal

mine

16.51 2.88 736.70 4.39

Bailongs

han coal

mine

16.51 2.65 736.70 4.14

Fifth

mine of

the

Yangmei

block

16.51 2.07 736.70 3.85

In addition, the outburst energy in Table 6 can be

converted into seismic magnitude, using the equation as

follows47

: 1.2 1.510 10 MW (17)

where M is the seismic magnitude. As shown in

Table 8, under the condition of tectonic coal, the

corresponding seismic magnitude of these 5 cases are

2.07 ~ 2.93. Generally, this seismic magnitude can not

be perceived or slightly perceived by human on the

ground. However, the corresponding seismic magnitude

are 3.85 ~ 4.39, when intact coal outburst occurs. For

this seismic magnitude, there is an obvious sensation on

the ground, such as sloshing or shifting of small objects.

However, according to in-site investigation, there is no

strong dynamic effect as M > 4 seismic, when outburst

occurs.


23

As mentioned above, the coal particle or pulverized,

with a size less than 1mm, is the guarantee for rapid

desorption of adsorbed gas, which widely exists in

outburst coal. However, the generation of these coal

particles produce a large amount of newly added surface

area, require a lot of crushing work. Based on the

estimation for 5 outburst cases, the stress required for

intact coal outburst is almost impossible to reach;

meanwhile, the dynamic effect of outburst does not as

stronger as M > 4 seismic, which can be produced by

intact coal outburst. Therefore, it can be concluded that

the occurrence of outburst is related to tectonic coal,

only the outburst energy requirement of tectonic coal

can be satisfied in current mining depth.

Conclusion 6.

Coal and gas outburst is a kind of micro-geological

disaster that is controlled obviously by tectonic

movement. In this work, combined with the statistics of

tectonic types and coal characteristics for outburst cases,

the relationship between outburst and tectonic coal is

obtained, after the comparing of the tectonic coal and

intact coal and summarizing of the outburst basic

characteristics. Then, this relationship is demonstrate

from an energy perspective. The main conclusions are as

follows:

1) Tectonic coal is a product of tectonic movement,

and the primary structure of tectonic coal is destroyed in

tectonic process. Moreover, the pore structure of

tectonic coal also undergoes corresponding changes,

which determine that gas adsorption is typically stronger

for tectonic coal and gas desorption rate of tectonic coal

is initially greater than intact coal. Tectonic coal shows

a low strength and weak cohesion properties, it is highly

broken even powdered after the disturbance of mining

activities. Due to these properties, tectonic coal has a

higher outburst risk than intact coal.

2) There is a close relationship between the tectonic

movement and outburst, especially for the large-scale

outburst. About 95.4% of 87 outbursts occurred in the

tectonic zone, such as fault, fold and heavily structured

zones. Based on statistical results, the strength of coal in

outburst cases is more conformed to the tectonic coal.

Moreover, tectonic coal has been observed in many coal

fields, while outbursts mainly occurred in these zones.

3) The coal particle or pulverized, with a size less

than 1mm, is the guarantee for rapid desorption of

adsorbed gas, which widely exists in outburst coal.

However, the generation of these coal particles produce

a large amount of newly added surface area, require a lot

of crushing work. Especially, the generation of

pulverized coals (d < 0.1mm) have the greatest impact

on the newly added surface area, and these pulverized

coals create 68.1% of the newly added surface area.

4) Both tectonic coal outburst and intact coal

outburst expend large amount of outburst energy for the

crushing and transport of coal. The crushing of coal

expends the vast majority of outburst energy, when

intact coal outburst occurs. Based on the estimation

results, the stress required for tectonic coal crushing is

16.51 MPa; but the stress required for intact coal

crushing reaches 736.70 MPa. Meanwhile, under the

condition of tectonic coal, the corresponding seismic

magnitude of 5 outburst cases are 2.07 ~ 2.93; however,

the corresponding seismic magnitude are 3.85 ~ 4.39,

when intact coal outburst occurs.

5) The ground stress in current mining depths maybe

satisfy the requirement for tectonic coal outburst, but it

is almost impossible to reach the requirement for intact

coal outburst. According to in-site investigation, the

dynamic effect of outburst does not as stronger as M > 4

seismic, which can be produced by intact coal outburst.

Therefore, only the outburst energy requirement of

tectonic coal can be satisfied in current mining depth.

Acknowledgments

The authors are grateful for the support from the

Fundamental Research Funds for the Central

Universities (2017XKZD01), and A Project Funded by

the Priority Academic Program Development of Jiangsu

Higher Education Institutions (PAPD).

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26

Comprehensive prevention and control technology of mine pressure bumping in fushun mining area Guojun Li, Jiquan Sheng

Fushun Mining Industry Group Co. Ltd., Fushun City, Liaoning Province, China

ABSTRACT

Mine pressure bumping is the dynamic phenomenon of sudden violent destruction which is caused by the transient

release of elastic deformation energy of rock and coal mass surrounding underground roadway or mining face, offten

accompanying with the phenomena such as coal and rock mass expel, blast and blare. Mine pressure bumping can

cause the roof subsidence, and both sides convergence, floor cracking, blowing up, and even reaching roof rock of the

roadway. The strong impact of bumping can also make workers' body jolts, collisions, extrusion and other direct

injuries, and even other major secondary disasters like gas and dust explosion due to the damage of the roadway,

equipment and ventilation facilities. This paper presents five statements on the occurrences rules and characters of

pressure bumping, its comprehensive prevention and control technology, and considering bumping prevention

engineering starting from the design source of the entire mining as a fundamental measure etc.

Key words:

Fushun mining area, Mine pressure bumping, Occurrences, Comprehensive prevention and control, Design source of

the entire mining

Occurrences of mine pressure bumping in 1.

Fushun mining area

In 1933 Fushun mining area had already occur

pressure bumping. When Laohutai Coal Mine mining

coal in -400m during 1970s, mine pressure bumping

began to occur and with the mining depth increasing, it

becomes more serious. Especially during the period

from 1993 to 2001, the frequent occurred bumping

imposed great difficulties to coal mining and roadway

supporting, which seriously threaten the safety of coal

production. On January 26, 2002, the ML3.7 bumping

took place in the air return way of 78002

comprehensively mechanized mining face is the

maximum magnitude bumping in Laohutai coal mine.

In 1997 the special department in charge of the

prevention technology management of mine pressure

bumping and coal and gas outburst was established,

which gradually normalized the bumping prevention

technology. The advanced prediction technologies, such

as electromagnetic radiation, the effective pressure

release method like high pressure water injection are

introduced in succession. And cooperative research and

technological breakthrough on bumping prevention was

carried out with relevant science and research institutes.

In 1998, for roadway support in extremely thick coal

seam, anchoring and wire netting combined with "U"

shaped steel retractable sheds were widely used. In order

to improve the strength of roadway support, the U

shaped steel sheds were changed into O shaped steel

sheds with double wires netting for fully enclosed

support. At the same time the type of steel shed

developed into heavier types, so U36 type steel shed is

generally adopted. Since 2005, the hydraulic support has

been introduced into the roadway support field. The

chock type roadway hydraulic support, the advanced

roadway (stride) hydraulic support, and the gate type

hydraulic support are developed and designed to be

installed and used in places with greater threat by mine

pressure bumping. The application further improves the

strength, the integrity and the stability of the roadway

support.

During over one hundred years of mining history in

Fushun, coal mining method has undergone five

transformations. In 1996, the Fully-mechanized top-

caving mining method replaced the hydraulic backfilling

blasting mining method in the inclined longwall face.

The reform of mining method not only greatly enhances

the support strength of working face, but also improves

the production environment of underground coal mining

face. It increases the mechanization level of coal mines,

lightens the heavy labor of workers, and raises labor

productivity and comprehensive economic benefits of

mine. The more important is that the reasonable mining

and driving layout simplified the production system and

reduced the number of underground workers. Thus the

favorable conditions for mine ventilation, gas control

and on-site safety management were created. Especially

important roles it took in the prevention to pressure

bumping. The accident of pressure bumping on the

working face was eliminated. It has greatly enhanced the

comprehensive ability of the mine to resist the disaster.

In 2006, the last top layer mining face 83001 finished.

Since then all the protective seam mining activities in

the east, middle and west three panels of Laohutai coal

mine have completed, so all the mining and driving

operation is carried out within the scope of protection,

and the threat degree of bumping is fundamentally

improved. The number of bumping accidents is

significantly reduced, and the impact strength and


27

damage of bumping are greatly reduced.

Table 1

A statistical table of destructive bumping accidents from 2000 to 2017 in Laohutai mine

Year Number ML1 ML2 ML3 Max ML Damaged

roadway(m)

2000 124 64 54 6 3.5 833

2001 239 126 98 15 3.6 2585

2002 136 62 56 18 3.7 1625

2003 88 43 40 5 3.3 1232

2004 61 22 32 7 3.4 1001

2005 42 18 20 4 3.4 907

2006 20 11 7 2 3.2 170

2007 6 3 2 1 3.2 100

2008 4 2 2 0 2.9 15

2009 1 0 1 0 2.3 15

2010 5 0 5 0 2.9 20

2011 1 0 1 0 2.4 10

2012 0

2013 2 1 1 0 2.2 20

2014 1 1.8 34

2015 0

2016 0

2017 0

The occurrence characteristics of pressure 2.

bumping in Fushun mining area

Laohutai coal mine started to mine in 1907 and is

located in the middle of Fushun mining area. The main

mining coal seam is an extremely thick coal seam with

complicated structure. The thickness is from 0.6m to

110.5m, averagely 58m. It is distributed in zonal and

asymmetrical syncline and trends nearly EW. The roof

of coal seam is averagely 194.08m thick oil shale that is

very east to fall in. The floor of coal seam is tuff with an

average thickness of 42.7m. It adopts the development

mode of Development of horizontal tunnel in inclined

shaft stages with fully mechanized top caving mining

method and whole caving method for goaf management.

The designed production capacity is 3 million t/a, and

the certificated production is 2.4 million t/a. There are

seven air intake shafts (6 inclined shafts, 1 vertical

shaft), and two air return inclined shafts respectively in

eastern and western part of the coal basin. At present it

has two production stages, which is divided into three

mining panels, namely the shallow re-mining panel (-

430m stage), the eastern panel and the middle panel of

the deep stage (-730m stage).

According to the experiences of bump prevention,

the regional distribution of bumping in Laohutai coal

mine has the following characteristics: 1. the geological

structure zone near the fault or syncline axis is a

frequent occurrence of bumping. 2. The deeper to mine,

the more pressure bumping accidents take place. 3. The

isolated coal pillar area is the seriously threatened areas

of bumping. 4. The dynamic phenomenon of pressure

bumping shows very strong when the first layer of fully

mechanized top caving face is mining. 5. The bumping

occurs frequently in fully mechanized top coal caving

face at the early stage of mining, close to the final

mining, or when the shape of gob area is approaching

"square". 6. Less bumping phenomena occur in the

second layer of mining face which is protected and

pressure released area.

The comprehensive prevention and control 3.

technology of pressure bumping in Fushun

mining area

In terms of prevention and control of bumping,

Fushun Mining Industry Co. Ltd. has invested more and

more and have extensively carried out joint research

cooperated with institutions, colleges and universities.

The "four in one" comprehensive prevention and control

measures gradually came into being, namely the

forecasting measure, the danger-relieving measure, the

effect testing measure and the individual’s safety protect


28

measure.

Forecasting of pressure bumping 3.1

The forecasting of pressure bumping risk includes

two parts: regional prediction and local prediction. The

regional prediction is based on the law of the occurrence

of bumping, the analysis results of micro-seismic

monitoring and the experiences. According to the degree

of bumping threat, production areas of the mine are

initially divided into four type of zones, namely strong,

medium, weak and no bumping threat zones. Local

prediction is a comprehensive use of drilling bits

method, pressure observation method, electromagnetic

radiation method, micro-seismic monitoring method and

other technology to predict the risk of bumping in

detailed mining or driving operations locations in the

bumping threaten zone.

(1)Drilling bits method

The drilling bits method is a method to distinguish

the risk of bumping by constructing bore with diameter

42mm in the coal seam and according to the variation

rule of discharged pulverized coal and related dynamic

effects. This method has many advantages, such as

convenient construction, simple equipment, less

investment and accurate prediction. But it cannot be

monitored in real time, and there is some personal

danger in the process of on-site testing. The critical

value of drilling bits is 4kg, that is, S > 4kg, has the risk

of pressure bumping, while S<4kg has no danger of

pressure bumping; and if there are obvious dynamic

phenomena in the drilling process, it is also identified

as a bumping hazard.

(2)Pressure observation method

Through the observations of the support resistance of

the working face, the advanced support resistance of two

roadways, the displacement and separation of roof and

floor of the roadway and so on, the main roof weighting

situations, the periodic weighting interval, sign of roof

weighting, and the the relationship between bumping

and mining operations could be understood, then, the

prediction of pressure bumping and large area weighting

of the mining face could be forecasted. The equipment

used in the mine pressure observation method is

relatively simple, easy to operate, and the cost of

equipment is relatively low. Relatively speaking, the

monitoring precision is relatively low, and the other

factors of the monitoring area can greatly influence the

precision, and the prediction timeliness is relatively

poor.(Fig1, 2 are the pictures of pressure observation

equipment)

Fig.1. Displacement monitoring equipment

Fig. 2. support resistance testing device

(3)Electromagnetic radiation method

Electromagnetic radiation is based on the law that

the coal and rock mass will produce electromagnetic

radiation and acoustic emission signals when it is loaded

with deformation and fracture. The greater the load

stress is, the greater the electromagnetic radiation /

acoustic emission intensity is. The larger the loading

rate is, the stronger the deformation and fracture process

is, the stronger the electromagnetic radiation / acoustic

emission signal is. The amplitude of electromagnetic

radiation and acoustic emission can reflect the stress

state and the accumulated elastic strain energy of coal

and rock mass. The intensity and amplitude of

electromagnetic radiation / acoustic emission mainly

reflect the loading degree of coal and rock mass and the

strength of deformation and fracture. The pulses number

of electromagnetic radiation / acoustic emission mainly

reflects the frequency of deformation and micro-fracture

of coal and rock mass. YDD16 portable acoustic electric

monitor (Fig 3) combines acoustic emission and ultra-

low frequency electromagnetic induction technology. It

is a portable, multi channel, multi signal and non-

contact directional monitoring instrument.

Fig.3. YDD16 portable acoustic electric monitor


29

(4)Micro-seismic monitoring method

The micro-seismic monitoring system of Laohutai

coal mine is a ARAMIS M/E microseismic monitoring

system introduced from Poland EMAG company in

August 2008(Fig.4). Using this system, real-time

monitoring of underground micro-seismic events can be

realized, and the location and vibration energy of the

source can be accurately calculated. Based on the

statistics of monitoring data, the causes and laws of

mine tremor are analyzed, and the development trend of

mine shock and bumping is predicted.

The prediction of bumping risk with micro-seismic

monitoring system is mainly based on the statistical

comparison method to predict macroscopically. By

recording the time, space, intensity distribution and

changes of micro-seismic activity, the changes of wave

velocity and spectrum are analyzed to find the change

trend of micro-seismic activity, and to predict the area

and intensity of the possible occurrence of micro-

seismic events.

Fig.4. Micro-seismic system schematic diagram

The prevention measure of pressure bumping 3.2

Measures for prevention and control of rock burst

include regional control measures and local control

measures, and regional control measures shall be taken

first. Measures such as mining protective layer,

reasonable mining area layout, and decreasing mining

intensity are adopted to reduce the stress concentration

degree. By means of high pressure water injection and

pressure relief drilling, the physical and mechanical

properties of the coal and rock mass near the mining

face and the roadway are changed, the capability of the

accumulating projectile and the velocity of the release

projectile are weakened.

(1) Mining of the protective layer

The main coal seam of Laohutai coal mine is very

thick with an average thickness of 58m. It has not only

the bumping risk, but also the high gas content.

Therefore, the top lamination is used as a protective

layer to exploit. The crustal stress and the bumping

energy of the coal seam are released, and at the same

time, it can also play a great role in discharging a large

amount of methane.

(2)Optimization of mining and driving layout

According to the elevation and occurrence of coal

seam, the deep stage of Laohutai coal mine are divided

into three fully mechanized mining areas, east, middle

and west mining areas. The fully mechanized top

caving mining method is adopted. The mining and

driving layout of "two mining faces, two driving faces

and one preparation face" is formed. In the three mining

areas of deep stage "one mining face, one driving face

and one preparation face” are kept, and in each mining

area only one working face is arranged to avoid mutual

disturbance.

(3)Reduce the strength of mining

Reducing the production capacity of mine and

mining face is a very effective measure to prevent and

control rock burst. The advancing speed of the coal

mining face of the mine is controlled at 40m/ months.

By taking measures to reduce the mining strength of the

mining face, the activity of the overlying strata of the

working face becomes slow, the pressure is released

slowly, and the stress concentration degree is obviously

relieved.

(4) High pressure water injection

Fully utilized the different periods of driving or

mining, high pressure water injection and static pressure

water injection are implemented. At first high pressure

water injection is conducted, and the effect of coal

cracking is achieved. Then, the natural pressure of mine

water supply system is used to carry out static pressure

water injection. The water injection pressure and water

injection time must be ensured, so that after water

injection, the water content will reach more than 4% or

the increase of moisture content will be more than 2%.

According to the parameters of coal body hardness, the

layout density of water injection hole is determined

reasonably, and the effective wetting range of water

injection hole is verified by drilling detection and so on.

The general spacing of water injection holes is not

greater than the diameter of effective wetting range, so

as to ensure the coverage range of water injection. The

distance between water injection holes in this mine is

generally less than 15m. The angle and depth of water

injection holes are determined according to the location

of water injection and the thickness of coal seam. ，The

pressure of high pressure water injection is more than

8MPa, and the static pressure water injection pressure is

over 1.5MPa. The 3D2-SZ type high pressure water

injection pump is used. The rated pressure is 17MPa.

The working pressure is generally 8.0-13.0MPa and the

flow rate is 135L/min.

When the tunnel is driving, the high pressure water

injection is carried out. The water injection hole is

constructed in front of the tunneling. The hole is 10-30m

deep, and the rubber sealing device is used to seal the

hole. The water injection holes in the two sides and the

roof are constructed just keep 15m distance back to the

excavated face. Each group of boreholes is set at every

15m distance, and the depth of the hole toward each side


30

is 15m. The depth of the top hole is determined

according to the coal thickness of the roof. After the

work face laneway is ready to be completed, the water

injection is carried out before the mining. The water

injection hole is constructed in the two gateways

towards the coal body. The depth of the drill hole is

determined according to the length of the working face,

so that the drill holes of the two gateways are

overlapped to each other and cover all the coal. In the

process of working face mining, the water injection is

carried out at the same time. High pressure water

injection is applied to drill holes in front of hydraulic

support and the borehole are drilled toward the roof of

coal seam. Each day when the top coal is caved at the

end of hydraulic support, the boreholes are drilled in

front of the support. Once the boreholes are finished,

they are sealed immediately and start to inject water. At

the top coal caving shift and the maintenance shift water

injection can be continuously conducted.

(5)Pressure relief drilling

The pressure relief hole is a controlling measure to

eliminate or reduce the danger of bumping by drilling

method. Releasing pressure drilling can release elastic

energy accumulated in coal and rock mass and eliminate

stress increasing area. Under the condition of high stress,

the elastic energy gathered in the coal seam will destroy

the coal around the borehole and release the energy and

eliminate the risk of bumping.

In the heading face with danger of bumping,

advanced pressure relief drilling is implemented. The

mode of production is in the way of "two digging and

one drill", that is, in a cycle of three shifts per day, we

use one shift to drill the pressure holes and other two

shifts for driving. 7-14 advanced pressure relief holes

are conducted every day. The boreholes are arranged in

upper and lower rows (Figure 5), and the depth of the

holes are 10-15m, in order to cover the tunneling

construction, and the driving work is always carried out

in the pressure relief range.

Fig.5. A schematic diagram of horizontal section of pressure relief drilling

Effect test of prevention and control measures 3.3

The bumping hazard area has to been tested after

regional prevention and control measures and local

control measures had been conducted. The test method

and index are the same as that of the prediction of the

risk of bumping. If the test results show that there is still

a risk of bumping or the prevention and control

measures are not put in place, it is necessary to

reformulate the design and further implement the

prevention and control measures. Until the prevention

and control measures are effective, after the inspection

is confirmed that no danger of impact occurs, they can

organize production under the condition of safety

precautions.

Safety protection measures for bumping 3.4

Safety protection measures are the last line of

prevention and control of bumping. At the same time, it

is also the lifeline of employees. Therefore, even if

regional and local measures are taken, the safety

measures cannot be neglected.

(1)Reinforcement of roadway support

The successful reform of supporting form fully

embodies the three characteristics of "large initial

bearing capacity, good shrinkage and strong support

strength", which creates favorable conditions for the

comprehensive popularization and application of fully

mechanized top caving technology, and provides a

strong guarantee for the treatment of hidden dangers

such as ventilation, especially the pressure bumping

problems. ①Composite support of U type steel shed (O type

shed) with anchor and wirenet

Composite support of U type steel shed (O type shed)

with anchor net（Fig.6），has the characteristics of

large initial setting force, high support strength, strong

impact resistance and complete supporting structure. It

realized the full section support of the roadway and the

control ability of the two sides and the floor is strong. It

can effectively reduce the convergence and the bottom

drum speed and reduce the amount of the maintenance

and construction of the roadway. When rock burst

occurs, it can effectively protect the roadway and protect

the safety of workers and equipment. Even if the impact

load is too large, the O type shed can be compressed

through the lap section and still maintain a complete

supporting structure, which can effectively restrain the

roof, the bottom drum, the destruction support and the


31

reverse equipment, which can avoid the gas

accumulation in the ventilation system caused by the

tunnel failure and reduce the degree of personal injury to

a worker.

Fig.6. O type shed

②Roadway hydraulic chock support

The roadway chock type hydraulic support is used to

improve the roadway support strength (Fig. 7) for

roadway fork, geological structure area and coal pillar

edge. Each chock hydraulic supports is composed of

four hydraulic props and high strength top beam and

bottom beam. It pays special attention to strengthening

the supporting strength of the roadway floor. It has a

very obvious effect on reducing the impact strength and

effectively protecting the roadway space. A number of

supports are placed adjacent to the roadway direction,

that is to say, a solid roadway space is formed (Fig. 8).

Fig.7. Fork hydraulic chock

Fig.8. Roadway chock support

③ Advanced roadway hydraulic support

The leading stress concentration areas at the two

ends of the roadway near the working face, or the

heading face are significantly affected by the dynamic

pressure, and the pressure from the roof and the floor are

large and easy to cause bump. The use of the advanced

roadway support (Fig. 9) can improve the supporting

strength of the roadway, slow down the deformation

speed of the roadway roof, effectively protect the

roadway, and prevent the bumping and roof accident.

Two advanced roadway supports stand side by side,

called one set. The two adjacent bases of one set of

supports are connected with jacks for support moving,

which can carry out self-moving support in all kinds of

mine roadways.

Fig.9. Advanced roadway support

④Hydraulic roadway support for bumping

prevention

When the roadway of a mining face is in the primary

coal body or adjacent to geological structure, there is a

danger of bumping. The hydraulic roadway support is

adopted. This kind of roadway is greatly disturbed by

excavation, shows strong mining pressure. The roof and

floor and two sides of the roadway converge very fast,

the maintenance work of the roadway is huge, and prone

to occur bumping. So such kind of roadway should

avoid rigid support, and should adopt high strength and

retractable enclosed support.

The hydraulic roadway support for bumping

prevention (Fig. 10) is composed of top beam, bottom

beam and hydraulic prop. The arch type top beam is

composed of 2 curved beams. The bottom beam is

composed of 2 bases and 1 middle joint. The connection

between the top beam and the bottom beam is the

hydraulic prop. The whole support is in gate shape, and

is used at the same time with O type shrunk steel shed.

It is arranged alternately and parallel with the O shed.

The hydraulic roadway support for bumping prevention

has the characteristics of large initial support force and

strong supporting ability, which realizes full section

support to the roadway, and has strong control ability to

the floor and both sides of the roadway. It can

effectively maintain roadway section, prevent roadway

floor heave, and protect the safety of workers and

equipment.

According to the requirements of bumping

prevention, the way of roadway support is reasonably

chosen at the 58008-3 working face of Laohutai coal

mine. When the working face is designed, the

prevention and control of bumping should be taken into

consideration. The design of the roadway support is

based on the idea of the heavy and strong support. When

driving the roadway, the permanent support strength

directly meets the requirements of bumping prevention.

The compound support in the air return way is


32

composed of anchor and netting, O type shed with U36

type steel, and the special hydraulic supports. The

installed number of ZT2 x 3200-18/35 type advanced

roadway support is 44 and the total number of ZP6400-

27/42G type bumping prevention hydraulic support is

41. By strengthening the roadway support, it not only

plays the role of passive protection, but also plays a

certain active role in the stability of the surrounding

rock, so that the roadway will not be deformed and

destroyed during the whole process of mine earthquake

vibration.

Fig.10. Gate type hydraulic support

(2)Individual protective equipment

In the bumping attacking, the main part of the

wounded is the brain, accounting for more than 90%,

followed by chest mechanical damage, including rib

fracture and so on, accounting for 60%. The damage of

internal organs is mainly lung, heart, stomach and so on,

accounting for 19%, followed by arms and legs, which

account for 19%. In order to protect the safety of the

workers, the Fushun Mining Group has developed

individual protective equipment for shock protection,

including shock helmets and shock resistant vests, and

has applied for patent. The guard plate of impact helmet

and anti shock vest (Fig. 11) is made of lightweight

chemical raw material, impact resistant ABS resin. The

space between the body of the shock proof helmet and

people’s head is filled with a buffering cushion made of

a thick buffer material, with a wide hatband with a

length to be adjusted freely, easy to wear, easy to

operate and not easy to fall off. When the collision

occurs, those protective equipments slow down the

impact of an object to the body and the body's head,

chest and viscera can be effectively protected, so as to

greatly reduce the personal injury caused by the impact

of the pressure bumping.

Fig.11. Individual protective equipment

(3)Shockproof house

Shockproof house is also an anti impact personal

protective device developed by Fushun mining group. It

is to select the special reinforcement support in the local

area and set up the shock absorbing and emergency

shelter facilities in the area of the pressure bumping. It is

suitable for the fixed guard posts in the dangerous

bumping area, and provides a safe environment for the

workers with long time operation.

(4)Self rescue system of compressed air

Self rescue system of compressed air consists of air

compressor, air supply pipe, nozzle, protective bag and

other parts. The air compressor produces high pressure

air through the pipeline to the working place. The

branch line pipe is connected with the compressed air

pipeline and the air supply nozzles are installed on the

branch line, and the air nozzle is set in the protective

bag (Fig. 12). When accidents such as coal and gas

outburst and rock burst occur, ventilation system are

damaged, gas and other harmful gases will threaten the

personal safety of operators. The workers can get into

the compressed air protection bag or open the mask of

compressed air self rescue box (Fig. 13), open the valve

for the air supply, and the high pressure air in the pipe

can provide fresh air for the safe haven.

Fig.12. compressed air protection bags

Fig.13. compressed air self rescue box

(5)Binding and fixing of object and equipment

When pressure bumping occurs, the equipment and

materials stored in the roadway will fly and subvert

because of the violent vibration. In that case, the body

impact and extrusion will cause physical injury to the

workers, which is also one of the main forms of pressure

bumping on casualties. In order to reduce the damage of

bumping, Laohutai mine has set up the binding rules of

the regional objects in the rock burst, and binding all the

objects stored in the bumping area to prevent its

movement. With all those measures well arranged, the


33

better effects have been achieved.

(6)Control of personnel in dangerous areas

In order to prevent casualties caused by rock burst,

we should reduce the number of underground operators.

The production process and labor organization should be

optimized to minimize the number of workers. It is

strictly prohibited to organize various personnel

intensive production operations in the dangerous areas.

The regulations for controlling the personnel in the

production shift will not allow anyone to enter that do

not belong to the mining units. In the production shift,

none is permitted to enter the bumping dangerous area.

CH4 and CO sensors are added to instead of inspector.

During the non production shift when someone needs to

enter the bumping area to operation, special measures

must be made and be approved by the chief engineer of

the mine. At the same time, the total number of

operations in the bumping area cannot exceed 9, and the

number of workers in each group shall not exceed 5.

The fundamental measure to prevent bumping 4.

According to the actual conditions of the mine, the

influence factors such as coal seam occurrence,

geological structure, geo-stress field, mechanical

properties of coal and rock, mining history, mining

disturbance and other bumping factors are mastered and

fully considered, and a reasonable production succession

planning is compiled to avoid the adverse effects of

local stress concentration on preparation and mining.

The design of bumping prevention should be formulated

at the same time with mining design, and incorporated

into the mining design. Priority should be given to the

regional prevention and control measures of rock burst

to prevent the shortage of design. If not to do so, a large

number of local pressure relief measures should be put

into the excavation project, which will not only waste

human resources and material resources, but also form

many potential hazards.

The upper part of the 58008-3 working face of the

Laohutai coal mine is 58008-2 working face and gab

area of 507, 508 and 509, and the lower part is the gob

area of 707, 708 and 709. Most of the mining areas are

in the protection range. Only the roadway ranged from

180~570m in the air return way belongs to blasting

mined area where the protection effect is not sufficient

and this part of roadway is also affected by the northern

F25 fault (see Fig. 14). So it is still in the bumping

threatened area.

The layout of the production system was optimized

in the early stage of the design of the working face.

According to experiences in the past the north roadway

should be designed as transport way. Due to the

bumping prevention factors, the design practice of the

transportation in the north roadway has been changed. In

this way, the large transportation equipment is arranged

in the south roadway, which is near the coal seam floor

without the danger of pressure bumping, and ensures

that there are no workers in the operation with the

danger of the bumping, and provides sufficient time and

space for the implementation of various bumping

prevention measures.

Fig.14. The layout plan of 58008-3 mining face


34

Fig.15. Sketch map of 58008-3 mining face

Conclusions 5.

(1) We must strengthen confidence and firmly

establish the idea of "bumping accident could be

controlled and prevented". We should strengthen the

organization and leadership, establish a strong

professional team, standardize and strictly implement

the comprehensive measures to prevent pressure

bumping. Although the pressure bumping cannot be

eliminated, but by taking preventive safety and

protection measures, we can weaken the impact of

bumping, control the physical injury caused by the

impact of pressure bumping accidents.

(2) Increasing the safety investment to improve the

level of technology and equipment on bumping

prevention, investing a large amount of high strength

support materials and enhance the ability to resist

impact, is a strong guarantee for the prevention and

control of rock burst.

(3) The bumping prevention work should start from

the source, give priority to the implementation of the

regional prevention and control measures, make full

consideration of the pressure bumping factors in the

mining design, and reasonably choose the opening way,

the roadway layout, the coal mining method, the mining

order, the advancing direction and the roadway support

process to avoid the bumping risk to the maximum

extent.

A brief introduction to the author:

Guojun Li (June 1964), born in Faku County,

Liaoning Province, China, graduated from the

Department of mining in China University of Mining

and Technology in July 1988. In June 2007, he received

a doctor’s degree in engineering, a professor senior

engineer, and a gainer of special subsidy from the State

Council. At present, 74 papers have been published, 8

publications and one book are published. He had won 4

of the first prize of the provincial and ministerial level

scientific and technological progress awards, and 9 of

the second awards, 4 of the third awards. Currently, the

chief engineer of Fushun Mining Industry Group Co.

Ltd.

Jiquan Sheng (July 1979), man, born in Kaiyuan

County, Liaoning Province, China ,graduated from

Heilongjiang Institute of Science and Technology in

July 2002, a master's degree, a senior engineer in mining,

and the chief engineer of Laohutai coal mine of Fushun

Mining Industry Group Co. Ltd, the director of the coal

mining Committee of the Fushun coal society, the safety

production technical expert of the Liaoning provincial

government, Member of the technical committee of

China coal Industry Association.


35

抚顺矿区冲击地压综合防治技术李国君盛继权

抚顺矿业集团有限责任公司，中国辽宁抚顺

摘要：

冲击地压是煤矿井巷或工作面周围煤岩体，由于弹性变形能的瞬时释放而产生的突然剧烈破坏的动力现象。本文从冲击地压在抚顺矿区的显现、特点、综合防治技术、抓住设计源头等五个方面进行了综合论述。

关键词：

抚顺矿区、冲击地压、显现、综合防治、设计源头

冲击地压是煤矿井巷或工作面周围煤岩体，由于弹性变形能的瞬时释放而产生的突然剧烈破坏的动力现象，常伴有煤岩体抛出、巨响及气浪等现象。冲击地压灾害造成巷道顶板下沉、两帮收敛、底板开裂鼓起甚至接顶，强烈冲击对作业人员身体造成颠簸、碰撞、挤压等直接伤害，甚至因巷道、设备及通风设施的损坏，引发瓦斯、煤尘燃爆等重大次生灾害。

冲击地压在抚顺矿区的显现 1.

抚顺矿区 1933 年就发生过冲击地压，老虎台矿在上世纪 70 年代开采深度 400m 时开始发生冲击地压，随着开采深度的不断增加，冲击地压灾害也变得严重，特别是在 1993～2001 年间，矿井冲击地压频繁发生，给工作面生产和巷道支护带来了极大的难度，严重影响矿井的安全生产。2002 年 1 月 26 日，78002 综放工作面回风顺槽发生 ML3.7 级冲击地压，是老虎台矿开采以来最大震级的冲击地压。

1997 年，老虎台矿成立专门机构两防办，负责矿井防治冲击地压和煤与瓦斯突出管理工作，使防冲工作逐步走向规范化。陆续引进电磁辐射等先进预测预报技术和高压注水等解危卸压方法，并与相关科研院所开展防冲方面的合作研究和技术攻关。

1998 年，在特厚煤层巷道支护中广泛采用锚网和“Ｕ”型钢可缩棚复合支护， “Ｕ”型支架由 U 型棚发展为 O 型棚铺双网全封闭支护，并逐步向重型化发展，现普遍采用Ｕ36 型棚，提高了巷道支护强度。2005 年，开始将液压支护引入到巷道支护领域，先后设计使用了垛式巷道液压支架、超前巷道（迈步）液压支架和防冲巷道（门式）液压支架，在受冲击地压威胁较大的地点安装使用，进一步提升了巷道支护强度和支护的整体性、稳定性。

在抚顺上百年的开采历史进程中采煤方法经历了五次变革，1996 年综放开采试验成功，炮采倾斜长壁水砂充填开采转变为综合机械化放顶煤采煤法。采煤方法的改革，不仅极大地增强了工作面的支护强度，改善了井下采煤工作面的生产作业环境，提升了矿井机械化水平，减轻了工人笨重的体力劳动，提高了劳动生产效率和矿井的综合经济效益。更主要的是由于采掘布局合理，简化了生产系统，精简

了井下作业人员，从而为矿井通风、瓦斯治理、等现场安全管理创造了有利条件，尤其是在冲击地压的防治上起到了关键作用，杜绝了工作面冲击地压事故，大大提升了矿井抵御灾害的综合能力。2006

年，最后一个首分层工作面 83001 回采结束以后，老虎矿深部东、中、西三个采区全部完成保护层开采，采掘作业在保护范围内进行，冲击地压威胁程度取得根本性好转。发生冲击地压的次数明显减少，冲击的强度和造成的破坏都大幅降低。

冲击地压在抚顺矿区显现的特点 2.

老虎台矿开采于 1907 年，位于抚顺煤田中部，主采煤层本层煤为结构复杂的特厚煤层，厚度 0.6～110.5m，平均厚度 58m。呈带状、不对称向斜赋存，近东西走向。煤层顶板为极易冒落的油母页岩，平均厚度 194.08m。底板为凝灰岩，平均厚度为 42.7m。采用斜竖井阶段水平大巷开拓方式，综合机械化放顶煤采煤法，全部垮落法管理采空区。矿井设计生产能力 300 万 t/a，核定生产能力 240 万 t/a。井田中央布置七条入风井（六条斜井、一条竖井），东、西两翼各布置两条回风斜井。现有二个生产水平，划分浅部复采区（-430m 水平）和深部东部、中部（-730m 水平）等三个采区。

根据以往的防治冲击地压工作经验总结，老虎台矿冲击地压区域分布具有以下特点：一是断层附近、向斜轴部等地质构造带，是冲击地压多发带；二是采深越大，冲击地压发生的次数越多；三是孤岛煤柱区是冲击地压严重区域；四是综放面一分层冲击地压显现强烈；五是综放面开采初期、采空区“见方”及接近终采时，冲击地压发生比较频繁。六是二分层受保护的卸压带发生冲击地压较少。

抚顺矿区冲击地压综合防治技术 3.

抚顺矿业集团有限责任公司及老虎台矿在冲击地压防治方面，不断加大投入，联合科研机构和大专院校，广泛开展冲击地压防治技术研究，形成了预测预报、解危措施、效果检验、安全防护“四位一体”的综合防治措施。


36

2000～2017 年矿井发生的破坏性冲击地压统计表

年份次数 1 级 2 级 3 级最大震级破坏巷道（m）

2000 124 64 54 6 3.5 级 833

2001 239 126 98 15 3.6 级 2585

2002 136 62 56 18 3.7 级 1625

2003 88 43 40 5 3.3 级 1232

2004 61 22 32 7 3.4 级 1001

2005 42 18 20 4 3.4 级 907

2006 20 11 7 2 3.2 级 170

2007 6 3 2 1 3.2 级 100

2008 4 2 2 0 2.9 级 15

2009 1 0 1 0 2.3 级 15

2010 5 0 5 0 2.9 级 20

2011 1 0 1 0 2.4 级 10

2012 0

2013 2 1 1 0 2.2 级 20

2014 1 1.8 级 34

2015 0

2016 0

2017 0

冲击地压预测预报 3.1

冲击地压危险性预测预报工作包括区域预测和局部预测两部分。区域预测是根据冲击地压发生规律和防治冲击地压工作经验及微震监测结果分析，初步将矿井各生产区域划分为强、中、弱和无冲击危险区域。局部预测是综合运用钻屑法、矿压观测法、电磁辐射法、微震监测法等技术手段，对冲击危险区域中有采掘作业的局部地点进行冲击危险性预测预报。

（1）钻屑法

钻屑法是通过在煤层中打直径 42mm 的钻孔，根据排出煤粉量的变化规律和有关动力效应，鉴别冲击危险的一种方法。钻屑法具有施工方便、设备简单、投入少、预测较准确等优点，但钻屑法不能做到实时监测，而且现场施工人员测试过程存在人身危险。钻屑量的临界值 4kg，即 S≥4kg 有冲击地压危险，S<4kg 则无冲击地压危险；如打钻过程中有明显动力现象，也认定为有冲击危险。

（2）矿压观测法

通过对工作面支架阻力、两巷超前支护阻力、巷道顶底板移近和顶板离层等情况的观测，了解采场来压程度、来压步距、来压征兆、冲击地压与采动的关系，并对冲击地压及采场大面积来压进行预测预报。矿压观测法采用的设备相对简单，便于人员操作，设备成本相对较低，但相对来说监测精度较低，受监测区域其他因素影响较大，且预测时效性相对较差。（图 1、2 为矿压观测设备实物图）

图1 顶底板动态仪

图2 液压支架工作阻力测定仪

（3）电磁辐射法

电磁辐射法是基于煤岩体受载变形破裂会产生电磁辐射和声发射信号的规律，当煤岩体受载应力越大时，电磁辐射/声发射强度就越大；加载速率越大，变形破裂过程越强烈，电磁辐射/声发射信号越强。电磁辐射和声发射幅值能够反映煤岩体所处的应力


37

状态和所积聚的弹性应变能，电磁辐射/声发射的强度（幅值）主要反映了煤岩体的受载程度及变形破裂强度，电磁辐射/声发射的脉冲数（频数）主要反映了煤岩体变形及微破裂的频次。YDD16 便携式声电监测仪（图 3）融合了声发射及超低频电磁感应技术，是一种便携式、多通道、多信号、非接触定向监测仪器。

图3 YDD16便携式声电监测仪

（4）微震监测法

老虎台矿微震监测系统是 2008 年 8 月自波兰EMAG 公司引进的 ARAMIS M/E 微震监测系统（图4）。使用该系统可以实现对井下微震事件的实时监测，并且能够精确计算震源位置和震动能量。通过对监测数据的统计，分析矿震发生的原因和规律，预测矿震及冲击地压的发展趋势。

利用微震监测进行冲击危险的预测预报，主要是采取统计对比的方法进行宏观上的预测。通过记录统计微震活动发生的时间、空间、强度分布及其变化，分析波速、频谱的变化情况，来找出微震活动的变化趋势，预测微震事件可能发生的区域和强度。

图4 微震系统示意图

冲击地压防治措施 3.2

冲击地压防治措施包括区域防治措施和局部防治措施，并优先采取区域防治措施。采取开采保护层、合理采区布局、降低开采强度等措施降低应力集中程度；采取高压注水、卸压钻孔等措施，改变采场及巷道附近煤岩体的物理力学性能，减弱积聚弹性能的能力和释放弹性能的速度。

（1）开采保护层

老虎台矿主采煤层为特厚煤层，平均厚度达 58m，不仅具有冲击危险，而且瓦斯含量高。因此，将顶分层作为保护层开采，在卸除地应力、释放冲击能量的同时，还可以起到大量排放瓦斯的作用。

（2）优化采掘布局

老虎台矿按煤层赋存标高及产状，将矿井深部划分东、中、西三个综采区，实行综合机械化放顶煤开采。矿井形成“两采两准一备”的采掘生产布局，深部三个采区保持“一个回采，一个准备，一个备用”，每个采区只布置一个工作面，避免相互扰动。

（3）降低开采强度

降低矿井和回采工作面生产能力是防治冲击地压一项极其有效的措施。老虎台矿回采工作面推进速度控制在 40m/月，通过采取降低回采工作面开采强度的措施，工作面上覆岩层活动变缓，矿压缓慢释放，应力集中程度明显缓解。

（4）高压注水

充分利用巷道掘进至工作面回采的各个时期，实施高压注水和静压注水。先进行高压注水，取得煤体致裂效果后，再利用矿井供水系统自然水压进行静压注水。保证注水压力和注水时间，使煤体注水后含水率达到 4%以上或含水率增值达到 2%以上。根据煤体硬度等参数合理确定注水孔布置密度，使用钻孔探测等方法验证注水孔有效湿润范围，一般注水孔间距不大于有效湿润范围直径，以确保注水覆盖范围，老虎台矿实施注水孔间距一般不大于15m。注水孔角度和深度根据注水位置和煤层厚度来确定，高压注水压力不低于 8MPa，静压注水压力不低于 1.5MPa。使用 3D2-SZ 型高压注水泵进行高压注水，额定压力 17MPa，工作压力一般在 8.0-

13.0MPa，流量 135L/min。

巷道掘进时实施边掘边注，在掘进工作面向掘进前方施工注水孔，孔深 10-30m，采用橡胶封孔器封孔，进行高压注水。在已掘巷道内滞后掘进头 15m

处，向两帮和顶板施工注水孔，每隔 15m 打一组，帮孔深度 15m，顶孔深度根据顶板煤厚确定。工作面巷道准备完成后实施采前注水，在两顺槽向回采煤体施工注水孔，钻孔深度根据工作面长度确定，确保两顺槽的钻孔相互搭接，覆盖全部回采煤体。工作面开采过程中实施边采边注，对工作面架前煤壁向顶板施工钻孔进行高压注水，每天放煤班打钻，并立即封孔注水，放煤班和检修班连续注水两个小班。

（5）卸压钻孔

卸压钻孔是利用钻孔的方法消除或减缓冲击地压危险的解危措施。卸压钻孔可以释放煤岩体中聚积的弹性能，消除应力增高区。利用高应力条件下，煤层中聚集的弹性能来破坏钻孔周围的煤体，使煤层卸压、释放能量，消除冲击危险。

在有冲击地压危险的掘进工作面，实施超前卸压钻孔，采取 “两掘一钻” 方式作业，即在每天三班作业的一循环中利用一个小班时间打卸压钻孔，两个班掘进。每天打 7-14 个超前卸压钻孔，上下两排布置（图 5），孔深为 10-15m，以此来掩护掘进施工，使掘进工作始终在卸压范围内进行。


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图5 卸压钻孔布置平剖面示意图

防治措施效果检验 3.3

对已经采取区域防治措施和局部防治措施的冲击危险区域，进行效果检验。检验方法和指标与冲击地压危险性预测预报相同，如检验结果显示仍有冲击危险或防治措施落实不到位，则需要重新制定设计，进一步实施防治措施。直到防治措施有效，经检验确认无冲击危险后，方可在采取安全防护措施的条件下组织生产。

冲击地压安全防护措施 3.4

安全防护措施是防治冲击地压的最后一道防线，同时也是发生冲击地压时员工的生命线，因此即便采取了区域、局部治理措施，也不能放松安全防护措施。

（1）加强巷道支护

支护形式的成功改革，充分体现了“较大的初承力、良好的可缩性和较强的支护强度”三大特点，为综合机械化放顶煤技术的全面推广应用创造了有利条件，同时为一通三防、特别是冲击地压等隐患治理提供了强有力的保证。 ① 锚网 U 型钢棚（O 型棚）复合支护

锚网 U 型钢棚（O 型棚）复合支护方法（图 6），具有初撑力大，支护强度高，抗冲击能力强，支护结构完整等特点。实现了对巷道的全断面支护，对两帮和底板的控制能力强，可以有效降低巷道两帮收敛和底鼓速度，减少巷道维修施工量。发生冲击地压时，可以有效保护巷道，保护作业人员和设备安全。即使冲击载荷过大时，O 型棚可以通过搭接部收缩来卸压，仍然保持完整的支护结构，有效遏制冲击地压造成巷道冒顶、底鼓、摧毁支架及颠翻设备等现象，可以避免因巷道破坏造成通风系统中断引起瓦斯积聚，减轻冲击地压对作业人员造成的人身伤害程度。

图6 O型棚支护

②巷道垛式液压支架支护

对于巷道岔口、构造区域、煤柱边缘等冲击危险性较大地点巷道，使用巷道垛式液压支架，提高巷道支护强度（图 7）。每组垛式液压支架由四根液压支柱与高强度的顶梁、底梁组成，尤其注重加强了对巷道底板的支护强度，对减缓冲击强度和有效保护巷道空间具有十分明显的效果。沿巷道方向将若干支架相邻摆放支撑，即形成牢固的巷道空间（图 8）。

图7 岔口垛式液压支架

图8 巷道垛式液压支架

③巷道超前液压支架支护

采煤工作面两端头和掘进工作面掌子头等超前应力集中区域，受动压影响显著，顶底板来压大，易发生冲击地压。使用巷道超前支护支架（图 9），可以提高巷道支护强度，减缓巷道顶板移近速率，有效保护巷道，防止冲击地压和顶板事故。巷道超前支护支架，两架并列组成一组，相邻的两个底座之间连接有移架千斤顶，可在各种矿井巷道内进行自移行走式支护。

④防冲巷道液压支架支护

采煤工作面顺槽处于原生煤体或邻近地质构造的巷道，有冲击地压危险，采用防冲巷道液压支架支护。这类巷道受采掘扰动大，矿压显现强烈，巷道顶底板及两帮收敛速度很快，巷道维修维护工程量巨大，且极易发生冲击地压。这类巷道应当杜绝刚性支护，采用高强度可缩封闭式支护。


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防冲巷道液压支架（图 10）由顶梁、底梁和液压支柱构成。拱型顶梁由 2 段弧形梁铰接组成，底梁由 2 个底座和 1 段中节铰接组成，顶梁和底梁间由液压支柱支撑。支架整体呈 O 型，与 O 型可缩钢棚同时使用，交替平行布置安装在 O 型棚棚空内。防冲巷道液压支架具有初撑力大、支护能力强等特点，实现了对巷道的全断面支护，对巷道底板和两帮控制能力强。能够有效维护巷道断面，防止巷道底鼓，保护作业人员和设备安全。

图9 超前巷道液压支架

图10 超前巷道液压支架

老虎台矿 58008-3 工作面根据防冲要求合理选择巷道支护方式，工作面设计时超前考虑防治冲击地压要求，巷道支护设计按照重型强支护的防冲理念，在巷道掘进时永久支护强度直接达到防冲要求。回风顺槽冲击地压区域巷道采用锚网 U36 型钢 O 型棚复合支护，并配合采用 ZT2×3200-18/35 型超前巷道液压支架和 ZP6400-27/42G 型防冲巷道液压支架加强支护。共计安设超前巷道液压支架 44 架，防冲巷道液压支架 41 架。通过加强巷道支护，既起到了被动防护作用，又对围岩稳定起到一定主动作用，使巷道在矿震整体震动过程中不发生变形和破坏。

（2）个体防护装备

发生冲击地压人员受伤的主要部位是脑部，占 90%

以上，其次是胸部的机械损坏，包括肋骨折断等占60%。而内部器官的损坏主要是肺、心、胃等占19%，再其次为上下肢的折断也占 19%。为保护作业人员人身安全，抚顺矿业集团公司研发了防冲击个体防护装备，包括防冲击头盔和防冲击背心，并申请了专利。

防冲击头盔和防冲击背心（图 11）的护板采用轻便的化工原料抗冲击 ABS 树脂制作，防冲击头盔内的盔体与人体头部之间夹填有较厚的缓冲材料制作的缓冲垫圈，有长度可随意调节的宽帽带，佩戴舒适，操作方便，不易脱落。可在发生冲撞时减缓物体对人体的冲击力，有效保护人体头部、胸部和内脏器官，从而大幅度降低因冲击地压带来的人身伤

害。

（3）防冲击屋

防冲击屋也是抚顺矿业集团公司研发的一种防冲击人身防护设施，是在冲击地压区域内选择局部地点特殊加强支护并设置减震和应急避灾设施，适用于冲击地压危险区域的固定性看守岗位，为长时间作业固定岗位作业人员提供一个配备防护措施的安全环境。

图11 个体防冲装备

（4）压风自救系统

压风自救系统由风泵、管路、供风嘴、防护袋等部分构成，风泵产生高压空气通过管路输送到作业地点，在压风管路接出支线管路并在支线管路上安装供风嘴，用防护袋将供风嘴套在袋中（图 12）。当发生煤与瓦斯突出及冲击地压等事故时，通风系统受到破坏，瓦斯等有毒有害气体会威胁作业人员人身安全。作业人员可以钻进压风防护袋内或打开箱式压风自救装置（图 13）的面罩，打开供风嘴阀门，压风管路内的高压空气可以为避险人员提供新鲜空气。

图12 袋式压风自救装置

图13 箱式压风自救装置

（5）物件设备捆绑固定

发生冲击地压时，巷道内存放的设备、材料等物件会因剧烈震动而飞扬、颠覆，对作业人员身体撞击、挤压造成人身伤害，这也是冲击地压造成伤亡事故的主要形式之一。为减小冲击地压危害，老虎台矿制定了冲击地压区域物件捆绑规定，对冲击地压区域巷道内存放的所有物件进行捆绑，防止其移


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动，取得了很好的效果。

（6）危险区域人员控制

为防止冲击地压造成集团伤亡事故，应减少井下作业人数。优化生产工艺和劳动组织，把作业人数降到最低，严禁在冲击危险区域组织各种人员密集型生产会战。冲击地压区域实行生产班禁员、非生产班限员的人员控制规定。生产班严禁任何人员进入冲击地压危险区域，增设甲烷和一氧化碳传感器代替瓦检员检查。非生产班需要进入冲击地压区域作业时，必须制定专项措施，报矿总工程师批准，同时进入冲击地压区域作业总人数不得超过 9 人，每组作业人数不得超过 5 人。

抓住设计源头是防治冲击地压根本措施 4.

依据矿井的实际情况，准确掌握并充分考虑煤层赋存、地质构造、地应力场、煤岩力学性质、开采历史、采掘扰动等冲击地压影响因素，编制合理的生产接续规划，避免局部应力集中给准备和开采带来不良影响。防冲设计要与开采设计同时制定，纳

入到开采设计当中。优先执行冲击地压的区域防治措施，防止设计上的先天不足，造成采掘工程开工后，还要投入大量的局部解危卸压措施，既浪费人力物力，还会因此形成许多现场隐患。

老虎台矿 58008-3 工作面上部为 58008-2 工作面已采区及 507、508、509 炮采已采区充填体，下部为 707、708、709 炮采已采区充填体。大部分开采区域均处于保护层保护范围内，只有北顺槽180~570m 受炮采区保护但保护效果不充分，处于原708 炮采区以北煤柱区，且受北部 F25 断层影响，存在冲击危险（见图 14）。

在工作面设计初期经过方案对比，对生产系统布局进行了优化，改变老虎台矿一直以来用北顺槽运输的设计惯例，使用北顺槽回风、南顺槽运输（见图 15）。这样使大型运输设备布置在贴近煤层底板没有冲击地压危险的南顺槽，保证有冲击地压危险的北顺槽生产班没有人员作业，并且为实施各项防冲措施提供充足的时间和空间。

图14 58008-3工作面布置平面示意图

图15 58008-3工作面布置剖面示意图


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结语 5.

(1)做好防冲工作必须首先坚定信心，牢固树立“冲击地压可治、冲击地压事故可防可控”思想理念。加强组织领导，建立强有力的专业队伍，规范并严格落实防治冲击地压综合措施，虽然不能杜绝冲击地压，但是可以通过采取防治措施减弱冲击地压强度，采取安全防护控制冲击地压事故造成的人身伤害。

(2) 加大安全投入提高防治冲击地压技术装备水平，大量投入高强度支护材料，增强抗冲击能力，是做好冲击地压防治工作的有力保障。

(3)防冲工作要从源头做起，优先执行区域防治措施，制定开采设计时充分考虑冲击地压因素，合理选择开拓方式、巷道布置、采煤方法、开采顺序、推进方向和巷道支护工艺，最大限度地规避冲击风

险。

作者简介：

李国君（1964 年 6 月—），男，辽宁省法库县人，1988 年 7 月毕业于中国矿业大学采矿系，2007 年 6

月获工学博士学位、教授级高工、享受国务院政府特殊津贴；截至目前已公开发表论文 74 篇、参与编写各类出版物 8 部、出版专著一部，获省部级以上科技进步奖项一等奖 4 项、二等奖 9 项、三等奖 4

项；现任抚顺矿业集团公司总工程师。

盛继权（1979 年 7 月—），男，辽宁省开原县人，2002.07 毕业于黑龙江科技学院，硕士学位，采矿高级工程师，现任抚顺矿业集团公司老虎台矿总工程师，抚顺市煤炭学会井工采矿委员会理事，辽宁省政府安全生产技术专家，中国煤炭工业协会技术委员会委员。


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Effects of gas pressure on the dynamic compressive strength of coal Z. Q. Yin, Z. Y. Chen, J. C. Chang, Z. X. HU, H. F. Ma, G. M. Zhao

School of Mining and Safety, Anhui University of Science and Technology, Huainan, Anhui, China

ABSTRACT

In consideration of the research requirements from gas-containing coal mechanics at deep coal mine, the stress states

of coal in deep-level projects is analyzed; and then there search topic of gas-containing coal under coupled gas-static-

dynamic loads is brought forward. A new testing system is modified by split Hopkinson pressure bar (SHPB), where

gas-containing coal specimen deforms and fails at medium to high strain rate and axial static preloading. With the

system, tests for gas-containing coal strength with four levels of gas pressures (0, 0.5, 1.0 and 1.5 MPa), and the

coupled loads tests are carried out under same strain rates. Findings exhibited the general law of stress-strain curves

and dynamic strength influence of gas pressure. The results show that, when the gas pressure is increase, the dynamic

compressive strength will decrease, but the average dynamic increase factor (DIF) of compressive strength increases.

These conclusions validate the functional accuracy and reliability of the testing using the device.

Key words:

Deep mine; Gas-containing coal, Coupled loads, Dynamic strength, DIF, SHPB


43

Optimal coal discharge of hydraulic cutting inside coal seams for stimulating gas production: a case study in pingmei coalfield

Baiquan Lin

School of Safety Engineering, China University of Mining & Technology, Xuzhou, Jiangsu 221116, China

ABSTRACT

The low permeability of coal seams is the primary difficulty for gas production. Hydraulic cutting is one of the most

effective technologies for increasing permeability and stimulating gas production. This paper investigates the

hydraulic cutting styles in boreholes and the optimal equivalent borehole diameter after discharging the pulverized

coal particles. This paper introduces the hydraulic cutting equipment and cutting styles and determines that the cutting

cylinder is the most suitable style for a soft coal seam that is prone to gas outburst. The numerical simulation suggests

that the radius of the stress-relaxed range increases quickly when the equivalent borehole radius is small, and the rate

slows when the borehole radius becomes large. Then the result was verified by a field experiment in Pingmei coalfield.

The gas drainage flowrate and stress-relaxed range greatly increased after initially outputting 2 tons of coal from the

borehole, whereas the second outputting 2 tons of coal did not appreciably increase the stress-relaxed range and the

flowrate. Too much or too little coal discharge is not optimal, and the borehole radius is recommended to be between

0.3 m and 0.4 m after the hydraulic cutting considering the pros and cons in Pingmei coalfield. The results are helpful

for the engineering application of the hydraulic cutting technology, for increasing coal seam permeability, and for

stimulating gas production. The method for finding the optimal coal discharge in this paper is also helpful for other

coal mines.

Key words:

Hydraulic cutting; Gas drainage; Stress-relaxed range; Gas outburst; Coal discharge


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Spatiotemporal multifractal characteristics of electromagnetic radiation in response to deep coal rock bursts

Enyuan Wang


ABSTRACT

Dynamic collapses of deeply mined coal rocks are severe threats to miners; in order to predict collapses more

accurately using electromagnetic radiation (EMR), we investigate the spatiotemporal multifractal characteristics and

formation mechanism of EMR induced by underground coal mining. Coal rock in the burst-prone zone often

exchanges materials (gas, water and coal) and energy with its environment and gradually transitions from its original

stable equilibrium structure to a nonequilibrium dissipative structure with implicit spatiotemporal complexity or

multifractal structures, resulting in temporal variation in multifractal EMR. The inherent law of EMR time series

during damage evolution was analyzed by using time-varying multifractal theory. Results show that the time-varying

multifractal characteristics of EMR are determined by damage evolution processes. Moreover, the dissipated energy

caused by the damage evolutions, such as crack propagation, fractal sliding and shearing, can be regarded as the

fingerprint of various EMR micro-mechanics. The dynamic spatiotemporal multifractal spectrum of EMR considers

both spatial (multiple fractures) and temporal (dynamic evolution) characteristics of coal rocks and records the

dynamic evolution processes of rock bursts. Thus, it can be used to evaluate the coal deformation and fracture process.

The study is of significance for us to understand the EMR mechanism in detail and to increase the accuracy of the

EMR method in forecasting dynamic disasters.

Key words:

Rock bursts; Electromagnetic radiation; Dynamic collapses; Electromagnetic radiation


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Numerical simulation study of hard-thick roof inducing rock burst in coal mine

Jiang He1, Linming Dou

1,2, Songwei Wang

1, Changhao Shan

1

1. School of Mines, China University of Mining and Technology, Xuzhou, Jiangsu 221116 China

2. Key Laboratory of Deep Coal Resource Mining, China University of Mining and Technology, Xuzhou, Jiangsu

221116 China

ABSTRACT

The numerical simulation study results show that, the mining induced seismic events occurring in the roof breaking act

on the mining surrounding rocks. Horizontal vibration of roadway surrounding particles will cause instant changes of

horizontal stress of roadway surrounding rocks; the horizontal displacement is directly related to the horizontal stress

but is not significantly correlated with the vertical stress; the increase of horizontal stress of roadway near surface

surrounding rocks and the release of elastic deformation energy of deep surrounding coal and rock mass are immanent

causes that lead to the impact instability of roadway surrounding rocks.

Key words:

Hard-thick roof, Rock burst, Numerical simulation, Horizontal stress; stress wave


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Static and dynamic loads superposition mechanism of coal-gas dynamic disaster and its prevention principles

Linming Dou1, 4

, Xueqiu He2, Ting Ren

3, Jiang He

1, 4, Zengyi Wang

1, 4

1. Key Laboratory of Deep Coal Resource Mining, Ministry of Education, China University of Mining and

Technology, Xuzhou Jiangsu 221116, China

2. School of Civil and Resource Engineering，University of Science and Technology Beijing, Beijing 100083, China;

3. University of Wollongong, NSW 2522, Australia

4. School of Mines, China University of Mining and Technology, Xuzhou Jiangsu 221116, China

ABSTRACT

Safe and orderly production of coal mines is seriously affected by dynamic disasters of gas bearing coal or rock. Based

on laboratory test, field test and theoretical analysis, static and dynamic loads superposition mechanism of dynamic

disaster are proposed, which are divided into high static load type, strong dynamic type and low critical stress type.

The expression containing static load, dynamic load and gas pressure is proposed. This paper summarizes the theory

and technology of monitoring and prevention of coal-gas dynamic disaster. The monitoring methods include partition

and hierarchical monitoring, electromagnetic emission technology, microseismical remote on-line monitoring system,

elastic CT and seismic CT technology, gas emission dynamic index technology. The prevention methods contain

intensity weakening theory, directional hydraulic fracturing technology and regional outburst prevention technology.

Prospects of future research about coal-gas dynamic disaster in China also involved. The mechanism of dynamic

disaster should focus on the quantitative study of gas migration in surrounding rock of roadway and stope, especially

the mechanism and prevention of dynamical disaster in super-kilometre deep well and major structural zone. The

monitoring and prevention of dynamic disaster should be based on the analysis of the dynamic disaster mechanism,

which contributes to construct the scientific monitoring and prevention system.

Key words:

Coal-gas dynamic disaster, Dynamic load, Static load, Gas pressure, Mine seismic, Monitoring and warning,

Prevention


47

Strong-Soft-Strong structural model for roadway preventing rock burst and its realizing technology

Mingshi Gao

School of Mines, Key Laboratory of Deep Coal Resource Mining, China University of Mining and Technology,

Xuzhou, Jiangsu 221116 China

ABSTRACT

Rock burst, which is usually collapsed, caved, and even closed in an instant as causing great harm to the mine safety

production. The reason of roadway surrounding rock failure induced by rock burst is that the elastic complementary

energy accumulated in the rock has been greater than the minimum kinetic energy of surrounding rock damage. The

greater the difference between the two value of the elastic complementary energy and the minimum kinetic energy is,

the more seriously the rock burst happen. According to the mechanism of roadway destruction by rock burst, a strong-

soft-strong (3S) structural model to prevent rock burst from taking place was established and analyzed. The result

shows that the 3S structural model has several characters, such as strength character of the structure itself, stress

adjustment, wave absorbing, deformation and anti-deformation, energy consuming, etc.. Stress criterion and energy

rule of the small inner support structure destroyed by rock burst were deduced. At last, based on the 3S model, the

engineering application was theoretically discussed; several techniques such as decreasing shock energy, setting

proper soft structure, and improving support strength were put forward to prevent roadway from being destructed by

rock burst. The soft structure realizing technique includes water-cracking technology, blasting-cracking technology

and gas-cracking technology.

Key words:

Strong-soft-strong structure, Rockburst prevention, Rockburst mechanism


48

Residual ground stress affected by the mining order and mining direction: a case study

Wei Yang


ABSTRACT

Ground stress is one of the major factors affecting coal and gas outburst. According to the statistical data, the average

dip angle of the maximum principal stress is 10.7 degrees, tending to be in the horizontal direction. Moreover, the

azimuth in the same geological unit is roughly the same. Based on the actual geological data and mining order of the

8th coal mine in Pingdingshan City in China, a numerical simulation model was established. The neighboring #1 and

#2 panels are in #1 coal seam. The surrounding stress distribution and evolution was studied when mining the two

neighboring panels. The result shows that the residual stress is greatly affected by the mining order and mining

direction. By optimizing them, the residual stress of the surrounding rock can be significantly reduced, so does the gas

outburst of the adjacent coal seam.

Mining #1 panel causes stress concentration in #2 panel, and mining #2 panel also compacts the gob of #1 panel. The

mining order in #1 coal seam greatly affects the gas drainage in #2 coal seam. There exists an optimal time for gas

extraction in the mining process. When mining in a coal seam with small dip angle, the mining direction significantly

affects the residual horizontal stress field. The residual stress of the surrounding rock can be significantly reduced

when the panel is mined along the direction of the maximum original principal stress. The difference among the three

principal stresses is also reduced vastly, which facilitates to decrease the gas outburst. The field data in the 8th coal

mine verifies the results. The #2 coal seam is 87.5 m below the #1 coal seam, and the outburst potential of #2 coal

seam is high. After the #1 coal seam is mined in the maximum horizontal direction, the outburst potential of #2 coal

seam is reduced to a large extent. The outburst in the protected range is significantly lower than that out of the

protected range when excavating a roadway. The results in this paper could be used as a valuable example for

exploitation far distance protective coal seam in high gassy coal mines.

Key words:

Gas outburst, Ground stress direction, Mining direction, Mining order, Protective coal seam


49

Regional clustering analysis of coal mine seismic events based on ant colony optimization algorithm

Xiaolin Li, Munezero Eugene, Zhiru Cai

School of Mines, China University of Mining & Technology, Xuzhou, Jiangsu 221116, China

ABSTRACT

The regional clustering analysis of seismic activity recorded in “Gucheng” coal mine in China are presented in this

paper. The analysis was conducted based on history time-space-strength data which is recorded and calculated by

using triaxial geophones monitoring system deployed in the mine zone. Clustering algorithm based on Ant Colony

Optimization was designed to group the seismic events efficiently. For each cluster, the evolution trend was analyzed

combined with geological conditions and mining activities to evaluate the relationship between seismic events and

regional stress distribution. The performance of the algorithm proposed was compared with typical clustering

algorithms including hierarchical clustering and K-means clustering, and the time-space correlation between clusters

and high energy seismic events were analyzed.

Key words:

Coal mine, Clustering algorithm, Seismic events, Ant colony optimization


50

Load distribution characteristics and anchoring force fluctuation mechanism of anchoring section under mining stress

Xigui Zheng

School of Mines, China University of Mining & Technology, Xuzhou, Jiangsu 221116, China

ABSTRACT

By analyzing the load distribution characteristics of the anchorage segment of the long anchor bolt with the resin

volume as the binder, the distribution of the bond stress distribution in the plastic zone of the anchorage section, the

elastoplastic zone and the elastic zone is studied. The evolution process of the adhesion and deterioration of the

anchorage zone under the action of the mining stress is further studied, the failure mode and the anchorage of the

second interface are analyzed. The response characteristics of the force jump are revealed, and the mechanism of the

roof bolt anchoring force step, the multi frequency oscillation and the complete loss of anchorage force are revealed.

The technical means of critical anchoring length, the rational anchoring factor, and the design method of the full

length prestressed anchorage and the project monitoring plan are put forward, which can be used for the long time of

the surrounding rock of the mining roadway. It provides scientific basis and theoretical support for stability and roof

safety.

Key words:

Mining stress; Prestress anchorage; Bond stress; Bond failure; Second interface.


51

Dynamic Early-warning of Rock Burst Based on Monitoring of Low-frequency Electromagnetic Radiation

Zhentang Liu


ABSTRACT

In order to study the response characteristic of electromagnetic radiation (EMR) signals with regard to coal rock

fractures, we conducted the monitoring experiment on EMR during the loading process of coal rocks. The relation

between EMR signals and paroxysmal fractures was analyzed from the perspective of EMR theory. Then time-series

extraction method of EMR representing coal rock fractures was proposed so as to further study the EMR's response

regularity and precursory features of instability during the deformation and fracture process of loaded coal rocks.

Deformation and fracture of loaded coal rocks is caused by the gradual accumulation of heterogeneously paroxysmal

fractures in coal rocks. The time series obtained by moving average method shows rather good correlation with

interior coal rock fractures, which has more obvious precursory characteristics of coal rock instability and dynamic

disasters. During the uniaxial compression process, coal rocks begin to fracture and low-frequency EMR signals are

more sensitive to local fracture with more abundant precursor information. Moreover, low-frequency EMR signals are

quite sensitive to small-scale dynamic phenomenon (coal burst) and the signal intensity can reveal its severity. EMR

signals representing coal rock fractures show an increasing trend as a whole and if the high signal lasts for some time,

it can be regarded as the precursor of rock burst

Key words:

Electromagnetic radiation, Coal rock fractures, Low-frequency, Forewarning


52

Research on the dual-porosity/dual-permeability solid-gas coupling model of moisture-containing fractured coal bed and its flow law Yuan Zhao, Yong Li

*, Shugang Cao

*, Bo Zhang, Hui Wang

State Key Laboratory of Coal Mine Disaster Dynamics and Control, College of Resource & Environmental Science,

Chongqing University, Chongqing 400044, P. R. China

ABSTRACT

Methane is one of the most serious disaster source in underground coal mine, and gas extraction is the most effective

method to prevention gas disaster. In this paper, the random fracture network has been built to represent the fracture

induced by tectonic stress and manual excavation disturbance. Moreover, the dual-porosity and dual-permeability

model was proposed by introducing the IM function of Comsol Multi-physics into random fracture network for

assigning value of matrix and fracture. Furthermore, the solid-gas coupling equations were embedded into Comsol

Multi-physics to simulation gas extraction. Through numerical computation of the proposed model, it can be

concluded that the porosity and permeability of matrix and fracture can be decreased with the increase of moisture

content and the influence of moisture content on fracture is much larger then on matrix. In addition, the increase of

moisture content can decrease gas extraction velocity, and then influence gas flow velocity distribution and gas

pressure distribution.

Key words:

Random fracture; Dual-porosity/dual-permeability; Solid-gas coupling; Moisture content

* Corresponding author. Tel: +86 188 8337 0816

E-mail address: [email protected], [email protected]


53

Study on precursors of rock instability: insight from acoustic emission (AE) Jing Li

*, Qianbing Zhang, Jian Zhao

Department of Civil Engineering, Monash University, Melbourne, VIC 3800, Australia

ABSTRACT

A better understanding of precursors of rock instability is significant for the prevention of underground dynamic

hazards such as the rock bursts. In this study, acoustic emission (AE) technique was employed to investigate rock

failure process under uniaxial loading. AE sensors (KRNBB-PC point contact sensor) were installed on the rock

sample to record AEs, and detected AE signals were digitized to 12 bits at a sampling frequency 2MHz by a high-

speed AE recording system. AE counts, frequency, and epicentre locations were analysed to show the evolution of

microcracks and early warning information of unstable failure. The experimental results show that three fracturing

stages could be distinguished according to AE counts, namely the microcracking initiation, stable slow development

and unstable rapid development. Besides, the change of AE frequency could indicate the precursors of rock instability.

At first, the AE frequency was relatively high which means that small-scale microcracks occurs gradually, while just

before strength peak, AE frequency tended to be lower indicating the rapid coalescence and development of unstable

large-scale fracturing. AE source localization makes a deeper insight into the microcracking evolution. During early

loading stage, a few acoustic emissions emerged and they were scattered within the rock sample. With the loading

increasing, more and more AE appeared and their distribution formed an obvious specific area, which was actually the

final rock failure plane.

Key words:

Acoustic emission; Microcracking initiation and development; Rock burst

* Corresponding author. Tel:+61 0452320917

E-mail address:[email protected]


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