Assessment of mining induced stress development over coal pillars during depillaring

International Journal of Rock Mechanics & Mining Sciences 48 (2011) 805–818

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International Journal ofRock Mechanics & Mining Sciences

1365-16

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Assessment of mining induced stress development over coal pillarsduring depillaring

A.K. Singh a, Rajendra Singh a,n, J. Maiti b, Rakesh Kumar a, P.K. Mandal a

a Central Institute of Mining and Fuel Research (CIMFR, under CSIR), Dhanbad 826001, Jharkhand, Indiab Indian Institute of Technology (IIT), Kharagpur 721302, West Bengal, India

a r t i c l e i n f o

Article history:

Received 29 June 2010

Received in revised form

29 January 2011

Accepted 15 April 2011Available online 8 May 2011

Keywords:

Bord and pillar

Field monitoring

Mining induced stress

Depth of cover

Roof caveability and stress meter

09/$ - see front matter & 2011 Elsevier Ltd. A

016/j.ijrmms.2011.04.004

esponding author. Present address: Camborn

, Penryn, Cornwall TR10 9EZ, UK. Tel.: þ44 1

4 1326 371859.

ail address: [email protected] (R. Singh).

a b s t r a c t

Earlier, an analysis of in situ observations at different sites of Indian coalfields was made to visualize the

development of mining induced stress over the coal pillars facing goaf line. An empirical relationship

was also attempted to estimate the range of influence and the value of ultimate induced stress (vertical)

over the coal pillars. However, the attempt was based on field monitoring data of only five depillaring

faces with varying geo-mining conditions. Considering a need of the Indian coal mining industry,

further field monitoring is done at 16 more depillaring faces with depth cover (average) range variation

from 44 to 244 m. The geo-mechanical properties of overlying roof strata of each site were also

determined to assess their caving characteristics in terms CMRI (now, CIMFR) caveability index, which

nearly varied from 1000 to 10,000 for the studied sites. Presenting a brief review of different studies

conducted for mining induced stress development, this paper discusses outcomes of the in situ studies

of mining induced stress development during depillaring under varying geo-mining conditions.

Considering the results of this study, earlier developed empirical relationship was, accordingly,

modified for estimation of the range of influence and the value of ultimate mining induced stress

(vertical) over the coal pillars.

& 2011 Elsevier Ltd. All rights reserved.

1. Introduction

Bord and pillars of different sizes and shapes are the two basicstructures associated with underground coal mining. Performanceof these two underground structures is responsible for the successof a mining operation below the ground. However, the perfor-mance of these structures is highly dependent upon two types ofstresses; mining induced stresses [1] and in situ stress [2]. For agiven site, the in situ stress is more or less static in nature but themining induced stresses over pillars/stooks keeps changing and ishighly influenced by the strata equilibrium dynamics during differ-ent stages of the underground coal mining activity. A familiar modelof the mining induced stress (vertical) development around anunderground mining face at different stages of working is shownin Fig. 1.

Many underground coal mines in India are operating at shallowcover, where board and pillar is the dominant mining method. Largenumber of coal seams has extensively been developed by formationof pillars to meet the increasing demand of coal in the country.Techno-economic scenario of the Indian coal mining industry

ll rights reserved.

e School of Mines, Cornwall

326 371839;

supported this strategy of coal production. It is reported that around3000 Mt of coal reserve is locked [3] in pillars under varying geo-mining conditions. Now, the industry is looking towards this hugeamount of locked-up coal in the pillars. However, undergroundextraction of these pillars is facing serious challenge due to presenceof difficult overlying strata. In general, underground coal mining inIndia often experiences strata control problems due to presence ofmassive and strong overlying strata [4–6]. Caving of roof strata is,generally, delayed and takes place after a large overhang during finalextraction (depillaring). The large overhang results in developmentof high value of mining induced stresses and dynamic loading ofsupports (both, natural and applied) during their breaking for fall.Here, an assessment of nature and amount of mining induced stressdevelopment is an important factor for proper pattern and designof supports to arrest adverse effects of the caving. Absence ofan estimation of the nature and amount of mining induced stressmay cause a substantial mismatch of the support during the finalextraction, which is a potential source of threat for safety ofunderground coal mining below competent roof strata.

Due to complex rock mass behavior under changing stressconditions of underground coal mining, an empirical formulationon the basis of field observations is, generally, adopted forassessment of nature and amount of mining induced stressdevelopment. Accordingly, CMRI (now CIMFR) earlier attemptedto develop an empirical formulation [1] and, therefore, undertook

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Width of excavation

Min

ing

indu

ced

stre

ss

Working horizon

Surface

Distressed zone

Mining induced stress development

Arch formation

Width of excavation

Stress level before mining



Surface subsidence

Width of excavation

Hanging cantilever

Min

ing

indu

ced

stre

ss

Angle of draw

Surface

Width of excavation Working horizon

Min

ing

indu

ced

stre

ss



Arch formation

Surface

Mining induced stress development Caved roof strata

Arch formation

Min

ing

indu

ced

stre

ss

Fig. 1. A conceptual model of mining induced stress (vertical) develeopment at different widths of excavations for an underground coal mining. (A) narrow working

without roof fall, (B) increased width of working with some roof fall, (C) wider working caused more roof fall without surface subsidence and (D) surface subsidence due to

further increased in working width.

A.K. Singh et al. / International Journal of Rock Mechanics & Mining Sciences 48 (2011) 805–818806

a field investigation to visualize the nature of development of themining induced stress under varying geo-mining conditions ofIndian coalfields. However, the study remained limited to onlyfive sites due to challenging nature of the associated fieldobservations. Recently, another Science and Technology (S&T)project is successfully completed [7], where sixteen more depil-laring faces were instrumented and monitored for assessment ofthe mining induced stress development. Nearly 150 vibratingwire stress meters were used for underground monitoring of thestress development with increase in dimension of the under-ground excavation due to pillar extraction. The quality of theoverlying roof strata of all these observed sites is assessedthrough geo-technical logging and laboratory testing of physico-mechanical properties of the freshly procured core samples.

In this paper, results of the above mentioned field studies arecompiled and an attempt is made to modify earlier formulations

[1] to assess ultimate value and range of influence of the mininginduced stress under varying geo-mining conditions of the coal-fields. Mining induced stresses are of two types: vertical andhorizontal. It is vertical mining induced stress which, generally,threats safety of the supports and studied under this project. Theword ‘‘mining induced stress’’ in this paper refers to only ‘‘verticalmining induced stress’’, which is discussed in this paper.

2. Mining induced stress

Existing natural state of stress equilibrium around a coal seamis disturbed by an opening formed due to underground extrac-tion of a part of the seam. The load of the overburden directlyabove the opening, previously carried by the coal, is transferredfrom immediate roof to surrounding pillars. An increase in the

A.K. Singh et al. / International Journal of Rock Mechanics & Mining Sciences 48 (2011) 805–818 807

width of the opening, generally, increases the value of the mininginduced stress and its range of influence over the surroundingpillars till the fall of the overlying strata. Before roof failure, theamount of the transferred overburden load due to an opening ismainly dependent upon its width and depth cover of the seam [8].In general, the surrounding pillars experience a maximumamount of mining induced stress just before the main fall ofthe roof.

2.1. Assessment approaches

In general, tributary area method is used to estimate the valueof mining induced stress around a symmetrical excavation withlow percentage of extraction. The scope of the tributary areamethod ends with high percentage of extraction and roof stratafailure. Once the strata breaks and acquires a new state ofequilibrium, an assessment of mining induced stress over coalpillars around the excavation becomes a challenging task. Here, itis difficult to correctly assess the amount of overhang, horizons ofbed separation and bulking of the caved strata. In fact, failure ofoverlying roof strata is mainly governed by the geology andstrength of the strata. Due to further dimensional increase ofthe opening, mining induced stress is created by the immediateroof strata cantilevering over the goaf area and their magnitudesdepend, mainly, on the length and thickness of the roof strata thatoverhang inside the goaf area.

In past, a number of attempts [9,10] utilizing, both, simulation[11,12] and field observations [13,14] were made to understandthe nature and amount of mining induced stress variation in andaround an underground excavation due to coal mining. In addi-tion to visualizing the nature of the mining induced stressredistribution, results of these studies were used to estimate thein situ strength of a coal pillar.

For a longwall face, experiencing bulking controlled caving, there-establishment of cover pressure distance [15] is characterizedwith a three parameter power function in which the independentvariables are depth, excavation height, bulking factor and com-pressive strength of the rock fragments. Yavuz [15] found that anincrease in mining height causes increase in bulking factor of thecaved rock piles resulting increase in cover pressure distance(Fig. 2). Efforts are also made to visualize the mining inducedstress development under jointed and unjointed rock mass [16].Even with all these measurements, there is lack of sufficientmining induced stress development data and, therefore, the dataof two entirely different geo-mining conditions [17,18] arecombined for numerical modeling [19].

0

100

200

300

400

500

600

0 50 100 150 200 250 300

Cover pressure distance (m)

Dep

th (

m)

h1 h2 h3 h4h1 h2 h3 h4h1 h2 h3 h4

Soft rock Medium strong rock Strong rock

Fig. 2. Effect of excavation height on the cover pressure distance (after [15]).

2.2. Mechanism of pillar loading

As soon as a pillar is formed in a coal seam, mining inducedstress develops over it, which, initially, remains confined over theedge of the pillar and its value stays small. Depillaring adopts anumber of manners of pillar extraction but all these manners;basically, reduce the size of pillars around the extraction lineresulting corresponding increase in width of the excavation. Anincrease in the width of the excavation results increase in value ofthe mining induced stress. Once the value of the induced stressexceeds the uniaxial compressive strength of the coal, some sidedeformation/spalling of pillar is observed and the position of thepeak value of the induced stress shifts inside the pillar. In fact, thearrest of increased value of mining induced stress is due to tri-axial state of the loading condition inside the pillar. The mostobvious sign of high value of mining induced stress at deepercover or under massive roof strata is spalling from the pillar/stooksurfaces. Further increase in the stress pushes the position of thepeak value of the induced stress further inside the pillar [20]which brings even core of the pillar under its influence beforefailure. Lunder and Pakalnis [21] and Fang and Harrison [22] havedescribed the progressive stages of degradation of a pillar underincreasing high value of stress.

2.3. Influencing parameters

Majumder and Chakrabarty [10] found that the mininginduced stress increases with decreasing seam thickness. Thiswas explained by the fact the reduction in seam thicknessgenerates a higher strain in coal adjacent to roadway. This inturn produces higher stress. On the basis of field measurementsand laboratory investigations on simulated models, Jayanthu et al.[23] also found that the maximum vertical stress over rib andstook decreases with increase in working height during depillar-ing. Observed maximum stress levels [23] over rib, stooks andpillars for different seam thickness and depth of cover are given inTable 1. Observations and evaluation of effect of floor benching onpillar stability [24] showed that increase in pillar height lowers itsstiffness resulting low pillar stress. However, all these observedinfluences of height of extraction on mining induced stress are notvery significant.

Field investigations [1] showed that the nature of developmentof mining induced stress over pillar/stook at different stages ofdepillaring, for a nearly flat coal seam, is influenced by differentparameters like depth of cover, characteristics of overlying strata,distance from face line, extraction height and goaf treatment. It isobserved that the core of a stook remains intact during caving ofweak roof strata but caving of massive roof may lead to overriding.

Table 1Maximum vertical stress over rib, stook and pillar for different seam thickness and

depth cover in the numerical models (after [23]).

Particulars Depth (m) Stress (MPa) for different height (he) of the workings

he¼3 m he¼5 m he¼7 m he¼9 m he¼11 m

Pillar 60 2.22 2.39 2.41 2.37 2.37

120 4.14 4.39 4.39 4.65 4.45

240 8.26 8.52 8.76 8.71 8.77

Stook 60 5.42 4.92 4.94 5.06 5.24

120 11.20 10.36 9.92 9.76 9.76

240 22.30 18.31 16.55 16.40 16.50

Rib 60 10.94 9.17 7.85 7.01 6.28

120 12.40 11.60 9.85 9.89 9.79

240 21.60 11.26 9.98 9.20 9.11

Just before failure

Min

ing

indu

ced

stre

ss (

vert

ical

) M

Pa

12

2 m

6m

Pillar (w/h=3)

Min

ing

indu

ced

stre

ss (

vert

ical

) M

Pa

2.0

19m

Only side spalling

2 m

Pillar (w/h=9.5)

9

6

3

0

1.5

1.0

0.5

0

Fig. 3. Observed profile of mining induced stress over two pillars of same depillaring panel having different w/h ratios and facing goaf line under similar strata movement

conditions.


An increase in width to height ratio (w/h) directly increasesstiffness and changes nature of post failure characteristic of a coalspecimen. Similarly, a pillar of higher w/h ratio arrests the roofmovement and its core remains intact even against a massive andstrong roof strata. While, on the other hand, a pillar/stook of lowvalue of w/h ratio around a depillaring face starts yielding/deforming,which allows, relatively, increased amount of overlying strata move-ment resulting higher value of the induced stress. Fig. 3 shows fieldobservations [25] of ultimate mining induced stress developed overtwo pillars, with w/h ratio 3 and 9.5, respectively, facing goaf line attwo different places in the same depillaring panel. These two pillars/stooks were located at center of the face line and encountered goafedge against a void of sufficient dimension to cause maximumamount of mining induced stress. This study is conducted in adepillaring panel with less than 80 m depth of cover and thecontinuity of measurement in space, especially in the squat pillar,remained quite discrete but could demonstrate the role of pillarstrength and stiffness over the development of mining induced stressat shallow cover. Therefore, it became important to select the size ofan instrumented pillar during the study, which could iteratively betackled by experience and mentioned in Section 5.

2.4. Significance of mining induced stress

Most of the underground coal mines in India practice bord andpillar method, mainly, due to the existing favorable [26] geo-technical conditions of the industry. In fact, competency of coaland rock masses supports this approach for the initial stage ofmining i.e., development of a coal seam. However, the final stageof mining i.e., depillaring encounters problems of strata controldue to competent overlying roof strata. Here, monitoring ofmining induced stress development is directly associated withthe safety of underground workings and even, sometimes; studyof nature of development of the horizontal value of the inducedstress is used [27] for prediction of major strata movement. Anestimation of amount and range of influence of mining inducedstress provides considerable help in optimizing, both, natural andapplied support [28]. In fact, safety factor of a pillar involves itsstrength and stress over the pillar. CIMFR has developed [29]empirical relationship among different geo-mining parameters toestimate pillar strength, which is given as

S¼ 0:27sch�0:36þH

1500:6þ

150

H

� �W

h�1

� �MPa ð1Þ

where S is the strength of pillar, sc is the compressive strength ofone inch cubes of coal (MPa), h is the extraction height (m), H isthe depth of cover (m) and W is the pillar width (m).

However, there is a lack of a reliable norm to estimate mininginduced stresses, in and around a depillaring face, which makes itdifficult to assess the safety factor. Further, in India, high capacityroof bolts and cable bolts are being used for final extraction ofcoal but, initially, the permission granting authority was againstapplication of these types of supports. In fact, Indian coal measureformations are of Lower Gondwana age, and are known fordelayed and violent caving. It was apprehended that furtherreinforcement of the massive roof strata by roof/cable bolting islikely to form a more dangerous combination as the caving of roofstrata will further be delayed. But the knowledge of pattern ofmining induced stress redistribution and its favorable interactionwith the reinforcement [30] around a depillaring face provided alogical explanation in adoption of such modern and productiveapproaches during depillaring.

For an effective capture of coal mine methane and to controlgas explosion during underground coal mining, it is importantto understand characteristics of gas (methane) flow and itsemission. These two parameters are largely dependent upon thehydraulic property (permeability) of the coal seam, whichbecomes dynamic under the influence of mining induced stressduring the final extraction of coal. Therefore it is significant tohave an idea of nature and amount of mining induced stressdevelopment in and around an underground coal face to analyzethe dynamic nature of coal permeability. Based on laboratory testdata and the simplified field conditions, an empirical relationshipbetween the horizontal coal permeability and the vertical stresshas been reported [31] to understand the dynamic nature of coalpermeability.

3. Important models

Analytical, empirical and numerical approaches have beenadopted in past to estimate the value and range of mining inducedstress over pillars during underground coal mining. Considering aninfinite, elastic, isotropic and homogeneous nature of coal measureformations, Salamon [32] derived following analytical equation forstress distribution at the edge of a longwall panel:

shðxÞ ¼xqffiffiffiffiffiffiffiffiffiffiffiffiffi

x2�L2p ðfor x4LÞ ð2Þ

where sh is the total stress, L is the half width of the panel, x is thedistance from the center of the panel and q is the virgin in situ stress,which is given as q¼gH. H is depth of cover and g is density of theoverburden.

Mark [33] used a concept (Fig. 4) of abutment angle (b) toestimate the abutment load during final extraction of coal, which

Laminated (t = 15) Overburden

Homogeneous Elastic Overburden

Empirical Abutment Stress

Distance from edge of panel (m)

Stre

ss (

MPa

) 40

50

605040302010

60

10

00

20

30

σ

-100 800

Fig. 5. Comparison of the longwall abutment stress computed from the homo-

geneous elastic model, the laminated model and the empirical formula (after [37]).

Ls

HB

P/2

P

Ls Hβ

H tanβ

Mined out panel

Fig. 4. The conceptualization of a side abutment load angle (after [33]).


is known as ‘‘Analysis of Longwall Pillar Stability’’ (ALPS) and isused widely for designing of pillars in longwall gate-roads. As perthis empirical approach, the measured distribution of inducedabutment stress (sf) follows following the equation:

sf ¼3Ls

ðDs�LÞ3ðDs�xÞ2 ð3Þ

where Ls is the total side abutment load and Ds is the maximumhorizontal extent of the abutment stress from the panel edge(x4L and xoDs).

However, the results of different field measurements in thelate 1990s [34,35] did not match with this concept of constantabutment angle. Heasley [36] considered homogeneous stratifica-tion of overburden to derive an analytical equation for inducedabutment stress (st), which is given as

stðxÞ ¼ qL

ffiffiffiffiffiffiffiffiffiffi2Es

ElM

re�

ffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffið2Es=ElMÞp

ðX�LÞ ð4Þ

where Es is the elastic modulus of the seam, E is the elasticmodulus of the overburden, M is the extraction thickness (forx4L) and l is the lamination constant, which is given as

l¼

ffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffiffit2

12ð1�n2Þ

sð5Þ

where t is the thickness of lamination and n is the Poisson ratio ofthe overburden.

Taking typical values of geometric and rock mass parameters,Heasley [37] computed abutment stresses at the edge of a long-wall panel with the help of Eqs. (2)–(4), which are shown in Fig. 5.To be consistent with other derivations during a comparison,Heasley shifted the origin (x) of ALPS from the edge of the panel tothe center of the panel. Plots of this figure show some disagree-ments among these three approaches, which are well discussedby Heasley. On the basis of experience of field studies of mininginduced stress in Indian coalfields, it is observed that the results ofabove mentioned [32] empirical formulations (Fig. 5) are similar tothose under week overlying strata [7] of Indian coalfields. The results

of analytical and numerical approaches (Fig. 5) are similar to thoseunder strong and massive overlying strata [7] of Indian coalfields.However, it is quite rational to consider the role of strata lamination[37] during analysis of mining induced stress development.

Poulsen [38] considered Load Transfer Distance (LTD) conceptin bord and pillar mining for estimating the lateral extent of thebase to the pressure arch between pillars [39]. Relationships weredeveloped to determine the value of LTD in terms of depth ofcover (H) after statistical analysis of 55 measurements [39] donein flat lying sedimentary deposits and is given as

LTD¼ 1� 10�4H2þ0:2701H ðmÞ ð6Þ

Considering this LTD concept, Poulsen [38] established follow-ing expression for estimating the average pillar stress in the zoneof influence (ZI), if excavation of area (ae) is made in the pillar’szone of influence:

Average pillar stress¼ rgHpð2LTDþWe=2Þ2=½pð2LTDþWe=2Þ2-ae�

ð7Þ

where We is the effective width of pillar, r is the average densityof roof rock mass and g is the acceleration due to gravity.

This is an important approach for automation of pillar loadestimation, but LTD remains valid for symmetrical extraction andmay not be suitable for a caving depillaring face. Size of a pillarfacing goaf line of a depillaring panel keeps deteriorating to acritical level and encounters, relatively, complex strata move-ment. A detailed study of interaction of different sizes of stooks/pillars with different types of overlying roof strata againstdifferent dimensions of the goaf would be an interesting investi-gation but is difficult to be conducted at real site. However, up tosome extent, this interaction is visualized through a laboratoryinvestigation conducted on simulated models.

4. Numerical modeling

A two dimensional finite difference based numerical modelingapproach is adopted to study the development of mining inducedstress over pillars/stooks of different size in and around a depillaringface. Available knowledge and experience of Mohr–Coloumb StrainHardening/Softening (MCSS) module of the Fast Lagrangian Analysisof Continua (FLAC) software [40] is the main basis for selection of thisapproach. Following two equations were used to estimate vertical(sv) and horizontal (sh) in situ stresses [2] during the simulation:

sv ¼ 0:025H ðMPaÞ ð8Þ

sh ¼ 2:4þ0:01H ð9Þ

Movement of overlying strata along the existing beddingplanes is one of the main reasons for development of mininginduced stress over pillars. Since FLAC is a continuum modelingapproach and therefore ‘‘interface’’ facility available with thispackage is used to simulate bedding planes in models. In strainsoftening approach, the shear strength (tsm) and friction angle(j0m) is estimated through Sheorey’s failure criterion for rockmasses [41] but this failure criterion is non-linear. Consideringthis fact, the value obtained by this failure criterion is slightlyadjusted for its use in the linear Mohr–Coulomb criterion.Accordingly, the value of shear strength is increased by 10% andthat of friction angle was reduced by 51 (Fig. 6).

The MCSS parameters other than the peak cohesion, frictionangle and dilation angle are also required to describe the rate ofcohesion and/or friction drop as a function of plastic strain in thepost-peak region. These MCSS parameters of rock mass arecalculated empirically by performing back analysis. Differentvalues of tsm, j0m and dilation angle for the corresponding


changing values of shear strains for MCSS modeling are given inTable 2 [40]. On the basis of a large number of Indian case studies,Murali Mohan et al. [42] found that the value given in Table 2suits well for the MCSS model.

Using the MCSS module of FLAC, different models are run withchanging values of strength and massiveness parameters (Fig. 7)of overlying strata to assess their effect on the amount and natureof mining induced stress over rib pillars of different sizes. The

Table 2Incorporated variation of different parameters in the Mohr–Coulomb strain-

hardening/softening model [after 40].

Shear strain Cohesion (MPa) Friction angle (deg) Dilation angle (deg)

0.000 1.1tsm j0m¼5 0

0.005 1.1tsm/5 j0m¼7.5 0

0.010 0.0 j0m¼10 0

0.050 0.0 j0m¼10 0

JOB TITLE : Behaviour of pillar loading condition durin

FLAC (Version 5.00)

LEGEND

8-Apr-10 17:34 step 25263 4.500E+01 <x< 1.900E+02 4.500E+01 <y< 1.200E+02

Grid plot

0 2E 1

Central Institute of Mining and Fuel Research, Dhanbad, India

0.60 0.800 1.0

Fig. 7. Deformation of discretized grid showing beha

Shaeorey criterion

Mohr-Coloumb criterion

σtm 0 σ

5°−φ

φ

smτ1.1

τ

τ

Fig. 6. Schematic diagram showing the non-linear Sheorey criterion as against the

linear Mohr–Coulomb criterion adopted in FLAC3D (after [42]).

material properties used for this simulation study are given inTable 3. Results of one such study for a 3 m thick coal seam, whichis developed on pillars to full height at 100 m depth of cover with22 m�22 m pillar size and gallery width of 4.0 m, are shown inFig. 8. Decrease of stress on rib pillars (Fig. 8) just after caving isdue to failure of rib pillars. For rib pillars of width more than10 m, observed stress values just before caving and just aftercaving were almost same (except at the edge), whereas aconsiderable drop in the value of mining induced stress (failureof the pillar) was observed after caving for pillars less than 10 mwidth. Considering some general values of rock mass parametersfor numerical simulation, this simple study could show the role ofpillar size for mining induced stress development. However,generally, different available testing procedures fail to generatethe required input parameters for simulation of an actual siteconditions. As per our practice, the estimation of simulationparameters for particular geo-mining conditions is done [43]with the help of available empirical formulations only.

5. Field study

Considering the importance of empirical formulations formining design, Singh et al. [1] conducted a field investigation tovisualize the development of mining induced stresses duringdepillaring in Indian coalfields. On the basis of this field study,the best fit equation for ultimate induced stress (Su) over a coal

g underground

00 1.200 1.400 1.600 1.800 (10^2)

1.100

1.000

0.900

0.800

0.700

0.600

0.500

(10^2)

vior of roof strata during depillaring operation.

Table 3Material properties used in simulation study.

Roof formations Young’s modulus,

E (GPa)

Density, r(g/cm3)

sc (MPa) RMR

Roof 2–5 1.9–2.25 35 40–70

Floor 5 2.25 35 60

Coal 2 1.4 25–30 40–50

Rib thickness - 14m-12

-9

-6

-3

00 2 4 6 8 10 12 14

Distance from the edge of the rib pillarfrom goaf side (m)

Ver

tica

l str

ess

(MP

a)

Just before caving Just after caving

Just before caving Just after caving Just before caving Just after caving

Just before caving Just after cavingJust before caving Just after caving

Just before caving Just after caving Just before caving Just after caving


-9

-6

-3

00 2 4 6 8 10 12 14 16V

erti

cal s

tres

s (M

Pa)

Just before caving Just after caving

Rib thickness - 6m-12-10-8-6-4-20

0 1 2 3 4 5 6

Ver

tica

l str

ess

(MP

a)

Rib thickness - 8m-12-10-8-6-4-20

0 1 2 3 4 5 6 7 8

Ver

tica

l str

ess

(MP

a)


-9

-6

-3

00 1 2 3 4 5 6 7 8 9 10V

erti

cal s

tres

s (M

Pa)


-9

-6

-3

00 2 4 6 8 10 12

Ver

tica

l str

ess

(MP

a)


-9

-6

-3

00 0.5 1 1.5 2

Distance from the edge of the rib pillar from goaf side (m)

Ver

tica

l str

ess

(MP

a)

Rib thickness - 4m-15-12-9-6-30

0 1 2 3 4

Disatance from the edge of the pillar from goaf side (m)

Ver

tica

l str

ess

(MP

a)






Fig. 8. Trend of variation of vertical mining induced stress developed on rib pillars just before and after caving, when rib pillars are near to the goaf edge.


pillar is given as

Su ¼ 0:0033Iþ0:059H�9:85MPa ð10Þ

where I is caveability index and given as

I¼slnt0:5

5ð11Þ

where s is the uniaxial compressive strength in kg/cm2, l is theaverage length of core in cm, t is the thickness of the strong bed inm, and the factor n has a value of 1.2 in the case of uniformlymassive rocks with a weighted average of RQD of 80% and above.In all other cases, n¼1.

The range of influence (R) ahead of a depillaring face may beestimated by the expression

R¼ 0:106Iþ0:1H�12:45m ð12Þ

These expressions were derived on the basis of field observa-tions of only five depillaring sites of Indian coalfields and there-fore availability of more field data may further improve thereliability of these relationships with possible modifications.

A field study, similar to [1], of development of mining inducedstress over pillars (with face advance) in and around depillaring(caving) operations was conducted [7] at different mines ofdifferent coalfields of the country. All these depillaring panels

Table 5Caveability index of roof formations of Johila seam of Nowrozabad East colliery.

Name of

formations

Thickness of

bed (m)

Avg. length of

core (cm)

Comp. strength

(MPa)

Cav.

index (I)

MGSST 2.42 20.2 8.09 698.5

FGSST 0.96 15.0 31.66 1243.2

MGSST 0.95 47.5 18.85 2616.6

Intercalation 1.92 27.4 23.25 2507.8

Shaly Coal 0.5 –

Intercalation 2.53 28.1 27.01 3436.0

Coal 0.06 5.0 –

Intercalation 2.38 18.3 19 1462.7

MGSST 0.4 20.0 –

Carb.Shale 2.39 26.6 31.46 3653.7

Coal 0.29 10.0 –

Shale 0.18 17.0 –

MGSST 3.59 25.6 15.01 2055.9

Intercalation 0.65 20.0 27.94 1238.9

FGSST 1.6 22.9 18.47 1488.2

Shale 0.27 13.0 –

Intercalation 0.37 29.0 22.08 1111.6

Shaly Coal 0.55 28.28

Intercalation 3 11.0 28.81 1421.8

Conglomerate 5 39.08

MGSST: medium grained sandstone; FGSST: fine grained sandstone.


adopted intermediate mechanization (drilling, blasting and load-ing by machine) and diagonal line of extraction (dip to rise),except Anjan Hill mine. A straight line of extraction by a fullymechanized mining system utilizing continuous miner and shut-tle car combination is adopted at Anjan Hill mine. Even monitor-ing of development of mining induced stress at this mine wasdone almost continuous in time with the help of a microprocessorbased data logger. In general, liquidation of a pillar adoptedsplitting and slicing, leaving 2 m thick rib against the goaf.However, for the study purpose, width to height (w/h) ratio ofthe first instrumented stook/rib in a panel was kept, at least, threeto provide a strength equal to five times of overlying rockpressure i.e., 5gH (g is generic unit weight). Size of the secondinstrumented stook/rib in the panel, to be left inside goaf, wasvaried in the same observation panel depending upon theperformance of the first instrumented rib/stook. On the basis ofthese experiences, size of the instrumented stooks/pillars duringthe experiment at different selected sites varied to bear 3gH to7gH stress.

Our earlier attempt [1] was based on study at only five sites ofdepillaring with caving, while this study of nearly 4 years of timespan covered 16 (Table 4) different depillaring panels of differentmines. Average depth of cover of these sites varied from 44 and244 m, where nearly 150 vibrating wire stress meters wereinstalled at different selected observation stations before com-mencement of depillaring in the panel. The selected sites for thisstudy were practically free from major geotechnical disturbancesand all the coal seams were nearly flat (except four at SRP-1, SRP-3A, SRP-3 and RK-8 mines). It is also seen that the selected panelswere wide enough (generally of super-critical nature) [44] toexperience complete caving of overlying rock strata. To reduce thebarrier effect, the observation sites were selected from the middlerow of the pillars and their positions were chosen in such a waythat they were expected to experience maximum abutmentloading during depillaring in the panel. Stress meters (vibratingwire type) were installed in a horizontal hole drilled across eachselected pillar. The position of a stress meter inside the pillar waschosen in such a way that they remain, generally, in the center ofthe stooks/ribs after splitting and stooking/slicing of the originalpillars. The depth of these stress meters inside the hole variedbetween 1 and 7 m depending upon the final size of the stooks/ribs to be left for the observation. Instrumented stations remainedstationary and the extraction face overtook all these stations withincrease in dimension of the excavation.

Maximum amount of mining induced stress is, generally,noticed during the first main fall. Therefore, the stress meters

Table 4Details of mines where field investigations were conducted for the study.

Name of colliery Name of area and

company

Average height of

working (m)

Gradien

the seam

Somna Hasdeo area, SECL 1.9 1 in 21.3

Rajnagar Hasdeo area, SECL 2.6 1 in 13.4

S. Jhimar Hasdeo area, SECL 2.3 1 in 37

Nowrozabad Johila area, SECL 3.5 1 in 7

Churcha West Baikunthpur area, SECL 3.0 1 in 13

Chirimiri (Bartunga Hill) Chirimiri area, SECL 12.5 1 in 182

Madhusudanpur Kajora area, ECL 7 1 in 18

Alkusa Kustore area, BCCL 6.7 1 in 10

SRP-1 incline Srirampur (P) area, SCCL 2.0 1 in 2.5

SRP-3A incline Srirampur (P) area, SCCL 6.0 1 in 3

SRP-3 incline colliery Srirampur (P) area, SCCL 1.8 1 in 3

RK-8 incline Srirampur (P) area, SCCL 1.8 1 in 4

GDK-8 incline RG2 area, SCCL 10.5 1 in 9

Anjan Hill Chirimiri area, SECL 3.9 1 in 30

GDK-5 incline RG1 area, SCCL 4.0 1 in 5.5

GDK-2 incline RG1 area, SCCL 1.6 1 in 4.5

were placed in the panel to pick up the stress change during thestrata equilibrium dynamics of first major fall. Further, the depthof cover of the depillaring face also affected the performance ofthe natural support because the pillars at deeper mines encoun-tered side spalling during depillaring. The depth of the stressmeter inside the selected pillar was adjusted accordingly inadvance during instrumentation. To observe the nature andamount of mining induced stress across the pillar, a number ofstress meters were placed at certain intervals inside the horizon-tally drilled hole across the pillar. However, after some experiencethis practice was terminated looking at the consumption of theinstrument and the stress meters were installed in such a way topick up the maximum amount of stress. All monitoring, exceptAnjan Hill mine, were done manually with the help of a read outunit and thus so many times it could not be possible to pick themaximum value of the mining induced stress during the dynamicloading. However, up to some extent, application of a number ofinstruments for this purpose and analysis of the obtained data byCombined-Instruments-Approach (CIA) [28] is attempted toaddress this issue.

t of Name of

seam

Nature of goaf treatment Average depth

cover (m)

Name of

panel

C Caving 77 S7

4A Caving 172.5 7

Jhagrakhand 4A Caving 48 S11

Johilla Caving 80 TE—14

V Caving 244 65LW

Zero Caving 91 K1

Kajora top Caving 40 K-12(B)

IX Caving 238 7

3B Caving 95 3BS4

1 Caving 102.5 1S2

3A Caving 73.5 3AS1/A

4 Caving 65 4S2/A

3 Caving 252.5 BG1/6

Zero Caving 93.5 C

4 Caving 44.3 4S/8

3A Caving 235 33C


The characteristics of roof rock mass of each of these sites wereevaluated with the help of core samples procured through bore holedrilling. Core samples of, generally, ten times of thickness of the coalseam were obtained from the roof strata at each site. During loggingof the procured core samples, length of each and every piece of corewas measured to find the average length of core of each roofformation. All the procured core samples were brought to thelaboratory and tested for their physico-mechanical properties, which

y = 56.857eR² = 0.9538

0

10

20

30

40

50

60

70

80

0 10 20 30 40 50 60 70 80

Ver

tica

l str

ess

(MP

a x0

.1)

Goaf edge distance (m)

GDK8 Colliery

y = 39.993eR² = 0.8186

0

20

40

60

80

100

120

Ver

tica

l str

ess

(MP

a x0

.1)

Goaf edge d

Anjan Hill

y = 16R² =

0

40

80

120

160

200

240

280

320

360

Ver

tica

l str

ess

(MP

a x0

.1)

Goaf edge

GDK2

y = 69.437eR² = 0.9349

01224364860728496

108120

Ver

tica

l str

ess

(MP

a x0

.1)


Somna Colliery

y = 70.996eR² = 0.8682

0

26

52

78

104

130

0

Ver

tical

stre

ss (

MP

a x0

.1)

Goaf edge d

Rajnagar C

y = 38.526eR² = 0.974

05

1015202530354045

0

Ver

tica

l str

ess

(MP

a x0

.1)


Nowrozabad Colliery

y = 47.542eR² = 0.783

03672

108144180216252288324360

0

Ver

tica

l str

ess

(MP

a x0

.1)

Goaf edge

Churcha West

y = 10.299eR² = 0.9035

0

3

6

9

12

15

Ver

tica

l str

ess

(MP

a x0

.1)

Madhusudanpur Colliery

y = 114.43eR² = 0.59

022446688

110132154176198220

0

Ver

tica

l str

ess

(MP

a x0

.1)

Goaf edge d

Alkusa

y = R

05

1015202530354045

Ver

tica

l str

ess

(MP

a x0

.1)

Goaf edge d

SRP3A

y = 31.541eR² = 0.9113

0

5

10

15

20

25

30

35

40

45

0 5 10 15 20 25 30 35 40 45

Ver

tica

l str

ess

(MP

a x0

.1)


SRP3 Colliery

5 10 15 20 25 30 10 20 30 40 50 60

10 20 30 40 50 60 70

0 10 20 30

80

0


10 20 30 40 50 60 70 80 90

40 80

14 28 42 56 7

0 5 10 15 20 2

0 5 10 15 20 25 3

Fig. 9. Variation of mining induced stress (vertical) with respect to face position (aft

coalfields along with an exponential fitting curve and equation to each plot.

provided the idea of the strongest bed. Caveability index of each roofformation is determined by using Eq. (11).

For a massive formation of overlying strata, caveability indexestimation of different stratum within the ten times thickness rangewas not difficult due to presence of limited number of laminates.However, in the case of highly laminated roof strata, it becamedifficult and unreasonable to consider the index of each layer. In thissituation, the weighted average of only five thicker-most layers was

y = 15.664eR² = 0.9548

0

2

4

6

8

10

12

14

16

18

Ver

tica

l str

ess

(MP

a x0

.1) GDK5 Colliery

y = 6.544eR² = 0.9708

0

1

2

3

4

5

6

7

0

Ver

tica

l str

ess

(MP

a x0

.1)


South Jhimar Colliery

y = 83.932eR² = 0.8793

020406080

100120140160

Ver

tica

l str

ess

(MP

a x0

.1)


Bartunga Hill Colliery

y = 13.054eR² = 0.8947

0

4

8

12

16

20

Ver

tica

l str

ess

(MP

a x0

.1) SRP1 Colliery

y = 19.619eR² = 0.9193

0

3

6

9

12

15

18

21

24

27

Ver

tica

l str

ess

(MP

a x0

.1)

RK8 Colliery

istance (m)

Colliery

5.09e 0.8877

distance (m)

Colliery

istance (m)

olliery

distance (m)

Colliery

12

istance (m)

Colliery

36.065e² = 0.8836

istance (m)

Colliery

0 35 40 45 50 55 60 5 10 15 20 25 30

0 10 20 30 40 50 60 70

Goaf edge distance (m)0 10 20 30 40 50 60 70

Goaf edge distance (m)0 10 20 30 40 50 60 70

40 50 60 70

80

120 160 200

0 84 98 112 126 140

5 30 35 40 45 50

0 35 40 45Goaf edge distance (m)

0 5 10 15 20 25 30 35 40 4550 55 60

er application of CIA [28] over the observed data) at different mines of different


considered for the analysis purposes. On the basis of experience ofcaveability study of different coalfields, it was also made mandatorythat the total thickness of these five layers must be more than five

0

5

10

15

20

25

30

0Ver

tica

l str

ess

(MP

a x

0.1)


Girmit Colliery

020406080

100120

0Ver

tica

l str

ess

(MP

a x

0.1)


East-katras Colliery

0

10

20

30

40

50

60

Ver

tica

l str

ess

(MP

a x

0.1)


Parascole Colliery

5 10 15 20 25 30 35

10 20 30 40 50 60

0 10 20 30 40 5

Fig. 10. Variation of mining induced stress (verti

Table 6Summary of results obtained through field observations.

Colliery Depth

(m)

Ultimate induced

stress (MPa)

Caveability

index

Range of

influence (m)

Somna 77 11.8 4433 55

Rajnagar 172.5 12.4 3190 60

S Jhimar 48 0.688 1194 20

Nowrozabad 80 3.7 2208 60

Churcha 244 36.2 9168 200

Bartunga Hill 91 14.6 4386 70

Mahusudanpur 40 1.3.7 1845 40

Alkusa 238 21.3 4000 130

SRP-1 95 1.6 2432 40

SRP-3A 102.5 4.1 2687 50

SRP3 73.5 3.9 3500 30

RK-8 65 2.5 4223 55

GDK-8 252.5 7.1 5102 65

Anjan 93.5 9.8 4181 60

GDK-5 44.3 1.6 2766 50

GDK-2 235 34.3 5847 70

times of the height of extraction otherwise more layers wereconsidered to fulfill this criterion. An example of caveability indexestimation of different strata for a laminated roof formation of Johilaseam, Nowrozabad colliery, SECL is given in Table 5. The final valueof caveability index of the Johila seam, Nowrozabad colliery came to2208.3 as per the above mentioned criterion. Similar exercise wasconducted for each studied site for estimation of representativevalue of caveability index.

6. Results and discussion

As envisaged, the observed value of mining induced stressincreased with decrease in its distance from the face position andthe variations observed at all the selected sites are shown inFig. 9. In general, maximum value of the stress is observed duringmain roof fall. The estimated values of caveability index (I) andobserved maximum values of induced stress (ultimate inducedstress, Su) at each observational site are shown in Table 6. Resultsof the previous six studies [1] of depillaring faces (caving) areshown in Fig. 10. Values of caveability index (I) and ultimateinduced stresses of these sites are given in Table 7. Even afterextension of connecting cables of the stress meters to a safe place,

0

20

40

60

80

100

120

140

0

Ver

tica

l str

ess

(MP

a x

0.1)


Lachhipur Colliery

y = 58.442e-0.212x

R² = 0.9069

0

10

20

30

40

50

60

70

0

Ver

tica

l str

ess

(MP

a x

0.1)

Gooaf edge distance (m)

Madhusudanpur colliery

0

20

40

60

80

100

Ver

tica

l str

ess

(MP

a x

0.1)


Govinda Colliery

10 20 30 40 50 60 70 80

10 20 30

0

0 10 20 30 40 50

cal) with respect to face position (after [1]).

Table 7Geo-mining indices of the sites studied earlier (after [1]).

Name of colliery Name of seam Average height of

working (m)

Average depth

cover (m)

Nature of goaf

treatment

Cavaebility

Index

Range of

influence (m)

Ultimate induced

stress (MPa)

Lachhipur Sonachora 3.4 109 Caving 4817 50 13.3

Girmit Rana 3.0 54 Caving 2531 20 1.66

Parascole Upper Kajora 4.5 60 Caving 3135 30 4.71

East-katras X seam 2.5 146 Caving 3598 40 10.1

Govinda Mid. Kotma 3.0 50 Caving 4512 45 7.08

y = 35.612e-0.05x

R² = 0.9484

0

10

20

30

40

50

60

70

0

Ver

tica

l str

ess

(MP

a x

0.1)

Distance from goaf edge (m)

H < 200m

10 20 30 40 50 60 70 80 90

Fig. 11. Mining induced stress variation during depillaring at shallow cover

(deptho200 m).

H > 200 m

y = 66.596e-0.016x

R² = 0.8331

0

40

80

120

160

200

0 25 50 75 100 125 150 175 200 225V

erti

cal s

tres

s (M

Pa

x 0.

1)


Fig. 12. Mining induced stress variation during depillaring at deeper cover

(depth4200 m).

R² = 0.790

0

100

200

300

400

0

Ult

imat

ein

duce

dst

ress

(MP

a x

0.1)

Depth (m)50 100 150 200 250 300

Fig. 13. Variation of ultimate induced stress with depth cover.


observations of most of the stress meters were challenging andremained incomplete after an encounter of the hazardous condi-tions in and around caving faces. This situation, more or less, iscommon for all the selected sites but the observations at GDK-8Incline became erratic and incomplete quite early and quickreplacement could not be done. Therefore, the stress observationsof GDK-8 Incline are not considered for further analysis.

Most of the selected observation sites were at shallow depth ofcover because, still majority of the pillar extraction practices inIndia, is situated in this zone only. On the basis of field experiences,the demarcation line for shallow and deeper mine is considered tobe 200 m. Accordingly, the obtained results are divided into twoparts. Averaged values of mining induced stress for all the mineswith depth less than 200 m are combined and presented together inFig. 11 while those for deeper mines (4200 m) are shown in Fig. 12.Although the nature of stress developments is broadly divided intotwo depth cover zones, it is important to note that the nature andamount of stress development (Figs. 9 and 10) is site specific.

6.1. Influencing parameters

Height of extraction during depillaring influences the bulkingcharacteristic of caved material and also affects the heights ofcaved and fractured zones of overlying strata inside the goaf. But,in field, it was difficult to study the influence of height ofextraction over the development of mining induced stress devel-opment. The operational mining parameters are, more or less,kept same for all the observed sites, it is only depth of cover andcharacteristic of overlying roof strata, mainly, influenced thenature and amount of the stress development.

6.1.1. Impact of depth cover

Depth cover has already been identified [15,45] as a significantparameter for the development of mining induced stresses. Depthof cover influences the in situ stress condition, depositional

compactness of the strata and geo-physical properties of rockmass. According to a well-established norm of Indian Coal MinesRegulations, the pillar size of Indian coal mines increases with theincrease of depth cover. It is observed that the roof strata of arelatively deeper mine, Alkusa (caveability index 4000) exerted21.2 MPa ultimate induced stress, while that for the roof strata ofSomna (having similar caveability index; 4433 but shallower) wasonly 11.8 MPa. The variation of observed ultimate induced stresswith depth of cover for above mentioned mines is shown inFig. 13, while Fig. 14 represents variation of range of influencewith depth of cover for these mines.

6.1.2. Impact of overlying strata

It is reported and observed that the quantitative and qualita-tive nature of the mining induced stress is dependent upon the

R² = 0.562

0

50

100

150

200

250

0

Ran

ge o

f in

flue

nce

(m)

Depth cover (m)

50 100 150 200 250 300

Fig. 14. Variation of range of influence with depth cover.

R² = 0.737

0

40

80

120

160

200

0

Ult

imat

e in

duce

d st

ress

(MP

a x

0.1)

Caveability index

2000 4000 6000 8000 10000

Fig. 15. Variation of ultimate induced stress with caveability index.

R² = 0.620

0

50

100

150

200

250

0 2000 4000 6000 8000 10000

Ran

ge o

f in

flue

nce

(m

)

Caveability index

Fig. 16. Variation of range of influence with caveability index.

0

20

40

60

80

100

120

140

160

Somna

Rajnagar

S Jhimar

Now

rozabad

Chirim

iri

Madhusudanpur

SRP-1

SRP-3A

SRP-3

RK

-8

Anjan H

ill

GD

K-5

Lachipur

Girm

it

Parascole

Eastkatras

Govinda

Ult

imat

e in

duce

d st

ress

(M

Pa

x 0.

1)

Observed value

Estimated value

Fig. 17. Comparison between observed and estimated values of ultimate induced

stress (deptho200 m).


properties of the overlying rock strata [1,15]. As per our practicalexperiences and different correlation studies, the characteristic ofoverlying strata is represented by caveability index (Eq. (11)).Field investigations conducted at Govinda and South Jhimarcolliery had nearly same depth of cover but their caveabilityindexes were observed to be 4512 and 1194, respectively. Thevalue of ultimate induced observed at Govinda mine is 7.08 MPawhile that at South Jhimar colliery is 0.69 MPa only. Even there isconsiderable difference in observed range of influences at thesetwo sites. The variation of ultimate induced stress and range ofinfluence with the caveability index for the observed mines areshown in Figs. 15 and 16, respectively.

6.2. Data analysis

Fifteen sets of new data of ultimate induced stress and range ofinfluence were obtained through the field investigations in andaround depillaring faces. Statistical correlation did not allowinclusion of results of three mines; Churcha West, Alkusa andGDK-2, probably due to higher depth of cover region. However,inclusion of results of the five previously studied sites [1] isaccepted and increased the amount of data. These data weresubjected to multi variant regression analysis to establish arelationship for the estimation of the ultimate induced stress

(Su) over a coal pillar against caved goaf. The adopted analysisresulted an expression for Su which can be written as

Su ¼ 0:025Hþ8:646� 10�4HI0:5 MPa ðfor Ho200mÞ ð13Þ

where H is the depth of cover and I is the caveability index.A comparison of the values of ultimate mining induced

stresses derived from Eq. (13) and those observed in the fieldare given in Fig. 17. This plot shows the largest disagreementbetween the two values for Chirimiri mine. In fact, depillaring atChirimiri mine experienced an entirely different type of stratamovement because the mining was done below a hill cap [46]with rapid change in depth of cover and the coal seam was placedabove the surrounding ground level. The observed values ofultimate induced stress at all the four mines; SRP-1, SRP-3A,SRP-3 and RK-8 are quite less than those of the estimated values,probably due to seam gradient reason.

Estimation of range of influence for application of advancesupport ahead of a depillaring face is done as per Sheorey’s model[17], which is based on depth of cover only. Above discussedresults suggest that the consideration of nature of roof strata anddepth of cover, both, may provide a better estimation. Accord-ingly, these two factors were considered for analysis resulting anexpression for the range of influence (R) as

R¼ 0:16Hþ9:63� 10�3Im ðfor Ho200mÞ ð14Þ

Fig. 18 presents a comparison of the values of range ofinfluence (R) derived from Eq. (14) and those observed in the field.

0

10

20

30

40

50

60

70

80

Somna

Rajnagar

S Jhimar

Now

rozabad

Chirim

iri

Madhusudanpur

SRP-1

SRP-3A

SRP-3

RK

-8

Anjan H

ill

GD

K-5

Lachipur

Girm

it

Parascole

Eastkatras

Govinda

Ran

ge o

f in

flue

nce

(m)

Observed value

Estimated value

Fig. 18. Comparison between observed and estimated values of range of influence

(deptho200 m).


7. Conclusions

The development of mining induced stress is observed to be a sitespecific phenomenon, which is being strongly influenced by thedepth of cover as well as nature of overlying strata. It is observedthat the role of monitoring of mining induced stress during depillar-ing under massive and strong roof is of considerable importance forsafety. Selection of number, place, range and pattern of instrumentsfor the monitoring is observed to be a challenging task due tocomplex interaction of roof strata with pillar/stooks in and around adepillaring face. Variation in geo-mining conditions of different sitesenhances the magnitude of this challenge. However, considering thepractical mining conditions of Indian coalfields, an attempt is madefor estimation of mining induced stress development ahead of adepillaring face of Indian coal mines at shallow cover. In fact, Indiancoal mining practices in last couple of decades have resulted lockingof considerably large amount of coal in pillars. The planned quantumjump in coal production strategy by Indian coal mining industry is,mainly, dependent upon the success of a pillar extraction process.The reported investigation and correlation of the obtained resultsmay be of some use for improvement in design and safety of theplanned underground pillar extraction program.

Acknowledgments

The authors are obliged to the Director, CIMFR, for hispermission to publish this paper. Amit Kumar Singh and SahendraRam, Senior Scientific Assistants, CIMFR provided considerablehelp in field work. The co-operation provided by the managementof different coal companies during the field study is thankfullyacknowledged. The study reported in this paper is based on aScience and Technology (S&T) project funded by the Ministry ofCoal (Government of India) and supported by Central MinePlanning and Design Institute Limited of Coal India Limited. Theviews expressed in the paper are those of the authors, and notnecessarily of the institute to which they belong.

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Documents

Assessment of mining induced stress development over coal pillars during depillaring