INDUCED CAVING TECHNIQUES AMENABLE TO UNDERGROUND COAL MINES FOR HARD ROOF MANAGEMENT

CHANDRANI D. PRASAD*, ACHYUTA KRISHNA GHOSH**

Abstract

Hard, massive, difficult-to-cave sandstone roof is commonly encountered in Indian coal measure during winning of coal. Though such roof ensures ground stability during development because of their high self-supporting capability, but often poses a serious techno-economic problem at the time of winning of coal due to its non-caving nature and high strain energy storing capacity. Mining with stowing or partial extraction may be technically feasible solutions, but in most cases, are not economically viable. In such condition, induced caving of roof holds the key of success of coal winning.

Of different induced caving techniques, blasting of roof in goaf either in underground or from surface, high-pressure water injection, hydro-fracturing, use of expanding materials (also referred as silent explosives), can be practised depending on their respective technical suitability and economic viability with respect to a given set of geo-mining conditions. However, induced caving by blasting is the only method practised in a few Indian collieries. While R&D is on to establish the high-pressure water injection system in India, this method is quite common in Chinese coal mines and hydraulic fracturing has been used successfully in some of the mines of Australia.

Successful application of these techniques depends upon a number of geomining factors such as depth of working, gassiness of seam, in situ stress distribution pattern, rock mass stiffness, porosity and permeability of strata, geological disturbances etc. and several legislative and technical parameters relevant to that technique, like diameter, depth and number of holes, type and amount of explosive to be used, etc. Hence, to select a techno-economically acceptable induced caving technique for any given set of geo-mining condition, systematic guidelines are required and a quick review should be made available. In this paper, an attempt has been made to present a quick review of all these techniques.

Introduction

Strata control is the major problem in underground coalmines where the immediate roof is massive sandstone. After accumulating good deal of strain energy, it fails all of a sudden, often with violence causing air blasts, rock bursts and/or coal bumps. Spalling of coal, collapse of faces and overriding of pillars are some other common phenomena associated with such roof that endangers the mine workings and sometimes result in heavy damages to supports, machineries and ventilation network, and serious and fatal accidents.

Several accidents have taken place in the past with huge loss of production and equipment. A well-known example is that of the longwall face collapse at Churcha West Colliery in 1990 and at Kottadih Colliery in 1997 in India, a major windblast in Moonee Colliery, Australia and huge caving in Majiliang Coal Mine, China that created a subsidence basin of 70000 m2 in area and 0.7 m in depth. An earth tremor of 3.2 had been detected. A schedule of windblasts occurs in Newstan Colliery between 1995 and 1997, a total of 23 events. In all the cases air velocities of range 40 m/s to 140 km/hr have been recorded. It was found that delayed

caving due to massive sandstone has created the trouble.

Extraction with stowing is the safest option, where the goaf is solidly packed with sand, but it is highly cost intensive and cumbersome. On the other hand, partial extraction, the other option where splitting is taken as final operation and small stooks are formed and left in situ, is low in production and productivity, and does not comply with the basic principles of mineral conservation because of very poor recovery. Yield pillar technique was developed based on the concept that if pillars with safety factor marginally left in situ over time under massive roof, the left over pillars/stooks allows the roof to cave slowly and gradually. However, the method has been banned in India since the roof failure in Parascole West Colliery.

However, a number of induced caving techniques viz., blasting of roof in goaf either from underground or from surface, high-pressure water injection, hydraulic fracturing, use of expanding materials (also referred as silent explosives) are available. Induced caving by underground blasting is the most popular method despite its limitations of amount of explosive to be used underground, risks associated with it, and ventilation problems. Large-diameter deep-hole surface blasting above uncaved goaf is suitable for shallow working depths, preferably not exceeding 100 m.

Hydraulic fracturing from surface can deal any depth but it requires knowledge of in situ stress. Sleeve fracturing in combination with hydraulic fracturing is suitable for fracturing any type of rock. High-pressure water injection is appropriate for weakening of porous and permeable roof rock to turn it cave-able.

Non-explosive expanding materials that have not been tried till date to break roof rocks in underground openings, but have been used for secondary breaking of boulders and winning small opencast benches in soft formations may be tried to induce caving.

Induced caving

All the induced caving techniques currently available in the global technological shelf are based on either of the two principles-

To break the immediate roof with good fragmentation and bulking factor so that it fills the goaf adequately to prevent overriding or air blast /windblast to any further sudden and massive collapse of higher roof; and

To weaken the roof, or in other words, to improve its cavability, by modifying its geotechnical characteristics by creating artificial planes of weakness in it and/or by reducing its physico-mechanical strength.

On the basis of these two principles, various available techniques are:

Deep hole surface blasting Caving by underground

blasting Hydraulic fracturing High pressure water injection Other possible techniques

like use of non-explosive chemical compounds

All these methods can be applied in two ways, pre-fracturing or post-fracturing and proper identification of strong beds that play dominant role in roof caving and loading at the face, is the prerequisite for successful application of any of above said techniques.

Pre-fracturing is a better choice as here roof rock is weakened in advance of working. When falling in goaf, an already fragmented or pre-fractured roof rock will further swell in volume making greater bulking factor to be achievable even in hard rocks (Choudhury, 2002). Unhindered advancement of face and reduction in level of periodic weightings on supports are other noteworthy advantages associated with pre-fracturing technique.

Alternatively, post-fracturing is the common way where caving is induced in the roof strata in goaf after winning of coal. This method requires less meticulous planning compared to pre-fracturing systems, but as the roof treatment operation is carried out in goaf, unlike pre-fracturing technique, it has limited control on the consequence in terms of rock fragmentation.

4.1 Deep hole surface blasting

In this method, a number of large diameter deep holes are drilled from the surface above the goaf area and explosive is placed in such a way as to ensure the breakage of the identified strong rock bed on blasting. If the identified rock strata are well above the coal seam being worked, non-permitted explosives may be used; otherwise only permitted explosives should be used.

An experiment was carried out at longwall panel P-1 (Fig. 1) of width 150 m, in Burhar VI B coal seam of thickness 2 to 3.33 m at Rajendra Project, SECL. Depth of coal seam was varying from 60 to 63 m. 4×503 T chock shield supports of were deployed to provide support resistance of 75 t/m2 at yield pressure of 40 MPa. The strong bed of high Rock Quality Designation (RQD) was found at a height of 17 to 20 m above the coal seam. Holes of 30 to 52 m depth, 6 to 12 in number, spaced at a distance of

45 to 75 m from the face were blasted using 350 to 1000 kg of Aquadyne explosive. Observation of leg closure on supports, subsidence advance due to weighting, convergence in the gate roads, cavity formation, spalling at the face and the weighting interval were made to study the severity of weightings. Average subsidence advance was 1.47 m to 1.95 m. No severe weighting was observed where goaf blasting was adopted. On the contrary, with few exceptions, severe weighting was observed where blasting was not practised.

Other than the risk of hazard from explosives, application of this technique is limited by depth from surface. Very deep holes may deviate and may not fulfil the desired purpose apart from increased cost of drilling. Also providing casing in weaker sections of rock is a difficult in very deep holes. This technique has also been tried in longwall projects of Balrampur, Kumda & Rajendra mines of SECL and roof weightings have been managed successfully (Bhati, 2003).

4.2 Underground blasting

It is the most commonly practised technique in Indian mines as no additional set up and laboratory tests are required for its application. A number of inclined holes towards goaf are drilled and charged in advance. These holes are then blasted out with the help of explosives after the goaf edge is reached. However, it has a number of limitations. It is not suitable for gassy mines due to several blasting hazards associated with the use of explosives, even a small crack filled with gas may cause explosion. Another factor is height of roof that can be dislodged is limited to 6-8m with the permitted amount of explosives which varies with degree of gassiness of seam

or any particular mine as shown in table 1. In addition to this, presence of fault planes may provide an escape route for explosion gases and thus reduction in utilised explosive energy effect. Also, drilling holes beyond 10 m length with usual coal drill is difficult.

Similar situation occurred in Churcha West Mine of SECL. Seam V having thickness of 3 to 3.45 m is being worked at a depth of 320 m with strata monitoring and induced caving by blasting. Pillars are of size 45 × 45 m with a gallery width of 4.2 m. The blasting pattern in practice is shown in Fig.2. Six to fifteen numbers of holes of 43 mm diameter drilled at an angle of 550 towards goaf of depth 7 m is blasted with P1 explosive. They are drilled in three rows such that after blasting they will provide a breakage plane along which the hanging roof will detach as in Figs. 2(a) & (b).

Cumulative and daily convergence at 76L/26X (15 m Dip) is shown in Fig.3. Both the cumulative and daily convergence increases due to weighting and get reduced after introduction of blasting on 15th, 20th, 23rd, 24th and 26th

Feb 2003 at a distance of 15-45 m and corresponding fall on 11th, 14th and 20th

Feb 2003. Though, severity of falls has been reduced, spalling of 0.5 to 1 m is observed at one pillar from pillar under extraction. Based on the instrumentation observations CMRI suggested blast to a height of 15 m for maintaining safe stress level in workings, but the mine management is finding it difficult to deal with and are managing with a depth of 7 m at present. Thus, it is necessary to develop existing induced caving techniques for greater depth.

4.3 High pressure water injection

It is mainly practised as a pre-fracturing technique, where the immediate roof is

broken in advance of mining. It is suitable for roof rock which loses its strength under water. From the results of petrographic test, water immersion test and permeability test are needed to determine the extent to which strength of rock reduces after absorption of water. Adhesion, absorption, hydration, wedge effect and dissolution are the mechanisms utilised to alter the mechanical properties of roof rock. Number of inclined holes towards goaf is drilled and water is injected at high pressure to fracture the roof rock after sealing 3-6 m length of hole by cement grout. The maximum injection pressure depends on the resistance of roof rock to passage of water, which in turn depends on permeability and aperture of joints, fissures and cracks. Quantity of water per hole and its duration of injection are other important factors. Rigorous efforts are going on in CMRI to establish this technique, as it is applicable to greater depth irrespective of gassiness of mine. In addition to that water infusion inside the rock covers a wider area and number of holes required is appreciably less than that of blasting. However, it may raise humidity of working areas at depths. Schematic arrangement for water injection is shown in fig.5.

A number of trials have been made at Churcha West Colliery and Parascole Colliery and results are quite encouraging. In 10W extended panel of Churcha West Colliery, 13 numbers of NX size holes of length 50-70 m were drilled at an angle of 25o–30o (Fig. 4). These holes were sealed to a length of 6 m by cement grout and water was injected at an initial pressure of 1400 Psi. The pressure was gradually reduced but rate of injection was increased. About 32,000 litres of water was injected at a rate of 68 l/min till water leakage from roof was observed. It was found that, in the water-injected zone, spalling and overriding of pillars

gets reduced significantly and roof fall occurred layer-by-layer accompanied by water (CMRI Report, 2000a). The total cumulative convergence also increased before periodic and local falls of the overlying roof rocks in the goaf (Fig. 6) this indicates a significant decrease in the dynamism and intensity of caving. While, India is still in developing stage, it is a well-known practice in Chinese coalfields. In working section No. 402, Majiliang Coal Mine, China the depth of seam no. 2 was 106 m. working height was 6 m, and the immediate roof was 4.5 m thick sandy conglomerate overlain by 50 to 100 m thick sandstone. The area of overhanging reached 151000 m2

before caving into the gob. When caving occurs the caved area of the roof in the gob area was approximately 125000 m2.

An earth tremor had been detected; the seismic shock measured was 3.2 in magnitude and 4 – 5 in violence on the surface. In consequence a subsidence basin was formed on the surface with a dimension of 70000 m2 in area and 0.7 m in depth. There were about tens of places severely fractured on the surface, and the largest width of crack was near about 4 m.

To reduce such overhangs high pressure water jetting was adopted. Boreholes were drilled into the roof area ahead of the face and water was injected at a high pressure. This method reduced the strength of the roof rock by creating various cracks and thus the cavability of immediate roof was enhanced. This method is a well-known practice now in Chinese coalfields (Linsheng, 1987)

4.4 Hydraulic fracturing

The basic principle of hydraulic fracturing is to isolate a section of borehole and create a fracture by applying hydraulic pressure [Fig. 7(a)] for determination of in situ stresses. The fracture so created propagates in a

direction normal to the minimum stress direction. This fracture can be utilised to induce caving in massive sandstone bed by drilling required number of holes and joining them by hydraulic pressure to provide a breakage plane. However, isolated part of a borehole should be free from geological weakness planes or pre-existing fractures. If the line of goaf is parallel to maximum stress direction an effective breakage plane can be obtained otherwise it will not help. In such case, serrata S-350 an instrument of sleeve fracturing technique, another in situ stress measurement technique, having semi-cylindrical friction shells with two gaps diametrically opposite surrounded by hard rubber membrane, can be utilised to create fracture [Fig. 7(b)]. It facilitates creation of fracture in any desired direction irrespective of field stresses. However, the cracks do not propagate far from the borehole. Thus a combination of this instrument and hydraulic fracturing technique can serve the purpose to a good extent.

Pure hydro fracturing has been tried in Kumda 7 & 8 inclines of Bisrampur area of SECL. Longwall retreating with caving using powered supports of about 500 T capacity is being practised at this mine. The working depth is having a depth of around 50-60 m with a face length of 150 m. Water was injected in boreholes at a depth of 28-29 m to weaken the stable block and subsequent to that light intensity fall started taking place and face condition remained normal (Bhati, 2003).

Similar technique adopted in Moonee Colliery, Australia enabled the mine management to regain the level of productivity after a temporary closure of mine there. Moonee Colliery is a coal mining operation in the Newcastle coalfields of New South Wales located in the Catherine Hill Bay area, which is approximately 100 km north of Sydney.

It operates on longwall extraction system. First longwall block commences in November 1997. Northern seam is a high volatile low sulphur medium ash coal, which is used for power generation. The seam was mined at a height of around 3.1 m of a total seam thickness of upto 4 m. In the roof the Booragul tuff member forms the immediate roof of the working seam, which was termed as claystone roof. Above this was the Teralba Conglomerate that was upto 40 m thick and composed of conglomerate, sandstone and some mudstone. The compressive strength of the conglomerate averages around 45 MPa and exhibits tensile strength around 8 MPa. Due to overlying claystone the development height is restricted top 2.8 m. A major windblast occurred on April 30 1999. Although there was no audible warning of impending roof fall and no micro-seismic warning of the event but the crew loader was blown approximately 3 m bodily and suffered multiple compound fractures of the left arm. The helmet he was wearing and the phone piece he was holding were blown more than 40 m through the adjacent cut-through. The mines department Inspector then issued a stop work notice. Later to reduce such overhangs hydraulic fracturing accompanied by a real time microseismic monitoring was introduced as a part of the production process at Moonee Colliery. This has enabled the mine to regain the level of productivity. Hydraulic fracturing has introduced consistent and far less variable falls, averaging around 5000 m2 (Peter, 2000).

Successful application of hydraulic fracturing alone requires the lithological knowledge of immediate roof strata as well as the in situ stress distribution pattern of the concerned region. Of course, it is a costly endeavour but will

not be so at greater depth. Because, days are not far off when mining industry has to opt for deeper mining that itself necessitates determination of in situ stress for much better and safe mine design, as shallow deposits are depleting very fast. Moreover, elimination of hazards associated with use of explosive is an added advantage.

4.5 Other Possible Approaches

Acconex is a compound developed by Associated Cement Co., Mumbai, which when mixed with potable water of 30% of its weight, expands and generates stress of 30 to 40 MPa in confined condition. This has been used to break large boulders as a replacement of secondary blasting, as well as tried in winning small opencast benches in soft formations, by pouring its mixture with right quantity of water into boreholes of suitable diameter and length. The mixture expands and exerts forces on the wall of the hole, thus initiating, widening and propagating cracks in surrounding rocks. However, the mixture is to be used within 10 minutes of its preparation and the depth of a hole should be 80 to 90% of the desired depth of breakage. The rate of reaction depends on environmental conditions, rock types, etc. and the effects begin to manifest after 10 to 12 hours.

This technique has not yet been tried for controlled fragmentation of in situ roof rocks whether in underground or from surface and is still at conceptual stage. However, the tests conducted at CMRI, Dhanbad on sandstone boulders indicate that it is possible to generate good breakage in all directions in high strength rocks by this process (Pal Roy, 1996). A sandstone boulder of compressive strength of 55 MPa and tensile strength of 4 MPa was drilled with holes of depth nearly 40% of the boulder width. Approximately 5 kg/tonne

of Acconex was used and the action started 18 hours after the pouring of the mixture and completed in 48 hours. The process seems to hold a good possibility to find applications in induced caving.

Selection Procedure

Of different options, stowing, particularly the hydraulic sand stowing version of it, is a proven and well-known one, especially with reference to control of subsidence. However, as this system is cumbersome, cost-intensive and hinders mining operation cycle reducing production and productivity of the total system, it is not a desirable solution for hard roof management nowadays. Constraints of availability of fill material (commonly river sand) and water sometimes restrict the application of this system. However, fly ash (particularly bottom ash) has been successfully tried to fill underground mine voids. Where fly ash is available in plenty, i.e. in the vicinity of thermal power plants, and water is adequately available, stowing may be a techno-economically viable proposition. In other cases, this option may be left beyond consideration.

Partial extraction of bord-&-pillar workings with splitting of pillars as the final operation, sometimes followed by partial or complete backfilling of galleries and splits, is also a proved method for subsidence control. However, for hard roof management this method should not be considered as it defies the basic principle of mineral conservation because of very poor (less than 15 %) recovery. Based on the review of the available techniques the selection of appropriate induced caving technique should be done in two steps of a) selection of technically feasible solutions and then b) the most economic alternative by comparing the costs of the selected methods if more than one choice is available.

Conclusion

In a considerable number of underground coal mines, roof rocks comprise of massive sandstone that does not cave regularly on its own in goaf during coal winning and causes high stress accumulation in roof and pillars. Finally, such a roof caves violently resulting in severe overriding and air blasts. Number of accidents has taken place in the past and still, after adopting induced caving techniques, weightings are experienced in working zones in many mines, though its frequency and magnitude has reduced. Improvement of the existing technique of underground blasting of goaf edge to make it more compatible to gassy conditions and to develop other options is the need of the hour. This idea is further supported by the facts that day by day coal mine workings are becoming deeper that itself necessitates determination of in situ stress for much better and safe mine design, but making backfilling more and more difficult and uneconomic. Elimination of hazards associated with use of explosive and other advantages definitely supports hydraulic fracturing. While each induced caving technique is having its own limitations, some of them are suitable for gassy mines and other for deep mines. It is obvious that any particular method may not be suitable for all the cases.

AcknowledgementThe authors are grateful to Director, CMRI, Dhanbad, for his permission to publish this paper. Views expressed are of authors only. The authors are thankful to Dr. G. Banerjee & Mr. A. K. Ray, Scientists, CMRI, Dhanbad for their support in preparation of this paper.

References

Banerjee, G., Ray, A.K., Singh, G.S.P & Yadava, K.P., Hard Roof Management – A Key for High

Productivity in Longwall Coal Mines, Journal of Mines, Metals & Fuels, Vol. 51, Nos. 7 & 8, July-August, 2003, pp. 238-244.

Bhati, N. P., Inducing the Caving by Hydro-fracture Technology, Journal of Coal Mining Technology and Management, Vol 8, No. 2, Feb 2003, pp. 1-9.

Choudhury, M., Pre-fracturing of Roof Rocks – A Possible Future Strategy for Successful Longwall mining in India, Journal of Mines, Metals & Fuels, Vol. 50, No. 6, June, 2002, pp. 219-224.

CMRI Report on Control of Massive and Hard Roof by Water Injection under High Pressure, Coal S&T Project, Central Mining Research Institute, Dhanbad, June 2000 pp. 16-40.

CMRI Report on Scientific Study Of Caving Characteristics and Associated Ground Control Problems in 57 LW & 65 LW Bord & Pillar Depillaring Panels in Churcha west Mine, SECL, Central Mining Research Institute, Dhanbad, April 2000 pp. 1-3.

Das, S. K., Modern Coal Mining technology, Lovely Prakasan, Dhanbad, 1994, pp 105

Peter H. & Flowers D., Risk Minimisation in Longwall Operations in Massive Goaf Conditions Using Microseismic and Hydraulic Fracturing Techniques, Australia, 2000, pp 1-3, www.qrc.org.au

Roy, P. P & Singh, R.B., Efficacy of “RAYDET”, “EXEL” and “Acconex” in Eco-Friendly Blasting Operations – Some Observations based on Experiments, Drilling & Blasting Editors - Pradhan & Ghose), 1996, MINETECH publications, Bhubaneswar, India, pp 71 - 73

Sarkar, S. K., Chatterjee, T. K. & Singh, S. K., Characterisation of Indian Coal Measures and Some Aspects of their Typical Behaviour, First National Conference on Ground Control in Mining, Editor - S. K. Sarkar, India, 1995, pp. 3 – 19.

Sheorey, P.R, Barat, D, Mukherjee, K. P, Prasad, R. K, Das, M. N, Banerjee, G & Das, K. K., Application of the Yield Pillar Technique for Successful Depillaring under stiff strata, Int. J. Rock Mech. Min. Sci. & Geomech, vol. 32, No. 7, 1995, pp. 669 - 708

Willian, J, Scoble, M & Pakalnis, V., Destressing Practice in Rockburst Prone Ground, Proceedings Fourth Conference on Ground Control in Mining, Editors – S. S. Peng & J. H. Kelley, West Virginia, July 22-24, 1985, pp 135-138

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Table 1 DGMS stipulations on maximum permissible charges in a shot holeType of

ExplosiveGassiness of Seam Winning Method

Max. permissible charge/shot hole

P1 Degree I Cut face 800P3 Degree I, II & III Cut face 1000P5 Degree I Blasting off-the-solid 1000P5 Degree II & III Blasting off-the-solid 565

Fig 1. Layout of Panel P – 1 at Rajendra Project, SECL

Solid side pillar

Goaf side pillar

1.6

B’B’1.2

1.0

A’

A’

Fig.2 Goaf edge blasting pattern for induced caving at Churcha west mine

1.8

Solid side pillarGoaf side pillar

αα

α

α = 450 - 550

1.61.01.2 1.0

Fig 2(a) Section B’B’

Floor

Roof

α α α

α = 450 - 550

1.81.81.8

Fig. 2 (b): Section A’A’

Floor

Roof

Date vs Convergence

05

10152025303540

12/1

9/2

002

12/2

3/2

002

12/2

7/2

002

12/3

1/2

002

1/4

/2003

1/8

/2003

1/1

2/2

003

1/1

6/2

003

1/2

0/2

003

1/2

4/2

003

1/2

8/2

003

2/1

/2003

2/5

/2003

2/9

/2003

2/1

3/2

003

2/1

7/2

003

2/2

1/2

003

2/2

5/2

003

3/1

/2003

3/5

/2003

Date

Co

nverg

en

ce (

mm

)Daily Conv.

Cum. Conv.

Fig.3 Cumulative and daily convergence in induced blasting zone at 76L/26 X (15 m Dip)

Fig. 4 Location of holes drilled for water injection in 10 W extended panel at Churcha West Mine, SECL

Fig 6 Cumulative convergence in water injected and non-injected zone

(a) (b)

Fig.7 (a): Basic step pressurisation of borehole in Hydraulic Fracturing

(b) Single fracture developed by S-350 system

P

Friction Shell

Fracture

Consolidation by compression

Fracture

Borehole

Fluid Pressure

Inflatable Packer

Fracture