Ph. D. INERIS Baroudi H Baucheron

DATA ANALYSIS FOR STRESS MEASUREMENTS BY OVERCORING: NEW OPTIMIZATION TECHNIQUES

The data processing for in situ measunements by overcoring (or stress-relief in gênerai) is classically performed by the least squares method. This appears well suited, because the relation y = F (a) between stress tensor (o) and displacements or strain (Y) is linear. In this case, the unknowns ait; stresses, and the known data of the problem are the measured displacements or strains and the elastic properties of the rock, those are preferably measured in-situ (biaxial test).

However if we wish to détermine the elastic properties (E matrix in the gênerai case) from the overcoring test itself, the previous relation becomes Y = F (o,E). This relation is non lineafrç to résolve it numerous optimization techniques are available. The optimization methods without gradient calculation seems well adapted for this problem.

INTRODUCTION

The knowledge of natural stress state in a rock mass is often an essential datum to solve most modelling problems in geomechanics. Among the quantitative methods of in situ stress measurements at great depth in a borehole, overcoring is usually preferred (fig. 1). Two différent types of measurements are obtained: radial displacements with the "U.S.B.M cell", and strains with the "CSIR triaxial strain cell" or the "CSIRO hollow inclusion gauge" which was developped in Australia (fig. 2).

The détermination of the stress tensor is done by assuming the rock as perfectly elastic and homogeneous. We obtain six components of the stress tensor expressed in a fixed référence sytem as follow :

(a) First of ail, détermine the elastic constants of the rock by means of loading and unloading tests on the core obtained from overcoring or by uniaxial compression tests on rock spécimens drilled out in the appropriate directions.

(b) Then, find the équations for the measurements (displacements or strains) in terms of the elastic constants and six stress components.

(c) Invert the équations by using the least squares method or similar methods.

After presenting the classical technique for data analysis we will discuss new numerical methods, that allow the détermination of elastic properties and stresses from overcoring data.

TECHNIQUES OF DATA ANALYSIS FOR STRESS MEASUREMENT

The détermination of the stresses with in situ overcoring measurement leads to résolve a linear System where the second member is the measurement vector (Y) and the unknowns are the six stress tensor components (a). The number of linear équations is usually greater than six, which involves that we can't inverse the matrix (A). We use in that case the least squares method wich consists of finding the solution which minimizes the residual error :

AX = Y with X estimator of a e = residual error = Y -AX e l e minimum o X = (A 1 A) 1 A 1 Y
This method gives many parameters which allows to estimate the quality of the linear régression such as the probability law of X, estimation of the stress tensor in a fixed référence System. But no information is given on the principal stresses because the relation between thèse, the six directions and shear stress tensor components is highly non-linear. In order to solve this problem, we used a statistical approach. The probability law of the principal stresses and their orientations is determined by using statistical simulations of the law given by the least squares method.

Other similar methods of data analysis for stress measurements like the weighted least squares or the residual values methods [START_REF] Chambon | Statistic analysis applied to rock stress measurements[END_REF] could be used but results are almost identical. The probalistic approach developped by Tarantola [8] and based on the maximum of likelihood gives interesting informations but the results must be carefully interpreted.

Kg. 2 : CSIRO cell

The coefficients of the matrix A dépend on the properties of the rock and the orientation of the hole. For spécial cases, isotropic bodies for example [10], there are known explicitly. In gênerai [START_REF] Amadeib | Rock anisotropy and the theory of stress measurements[END_REF], there are determined from Computing. Also, A is a function of the elastic constant vector E.

NEW FORMULATION OF THE PROBLEM

The theory presented hère is based on linear elastic behaviour. We suppose that elastic Usually, three vectorial norms are used :

X II = 1/2 x = Max x.
The dimension of the vector E can be 2,5 or 9 if weconsider the rock as isotropic, transversaly isotropic or orthotropic. Also the total number of unknorwns could vary between 8 and 15.

ALGORITHMS OF OPTIMIZATION USED

The classical techniques of optimization using gradients and hessian matrix cannot be uised in our case. In gênerai we hâve no information on the deri vability of the f miction f , that is why we used two algorithms for unconstrained minimization of a function of several variables that does not require the évaluation of partial derivatives : POWELL's and MIFFLIN's algorithm (see Appendix A and B for more détails on the methods).

Like many methods of optimization, thèse algorithms must be initialized. For any results, we must verify that the minimum obtained is not relative and that we hâve really f (E,a) = 0. In our case, there is not only one solution in gênerai. The results given by the two methods are the nearest of the initial values. The Powell's algorithm [7] is easy to iinplement. It can be link with the family of methods using conjugate directions. The basic idea is to find the minimum of the function successively in n conjugates directions (n is the number of variables). This method requires at each itération n one variable minimization. For this we used a golden section method [START_REF] Brent R | Algorithms for minimization without derivatives[END_REF][START_REF] Minoux | Programmation mathématique : Théories et algorithmes[END_REF]. The Mifflin's algorithm [START_REF] Mifflin R | A superlinearly convergent algorithm for minimization without évaluation dérivâtes -Mathematical Programming[END_REF] is a second order extension of the method of local variations and it does not require any exact one variable minimization. The method retains the local variations property of accumulation points. This extension makes the algorithm an approximate Newton method and its convergence is in gênerai super-linear.

The efficiency in the convergence of the two methods is measured by the number of évaluations of the function F. Mifflin needs in gênerai fewer calculations than Powell. For the same initial values, the results of the algorithms are identical except when there are several solutions in a limited space. In this case, Mifflin gives the resuit which is the nearest to the initial solution while Powell give a resuit which could be différent.

APPLICATION

We présent hère an example of use of thèse two methods. We consider the rock to be isotropic. We used data from the CSIRO HI cell for which the values of the elastic constants and the components of the stress tensor were known. The tensor corresponds to the stress state existing at a deptli of 450 m. The minor principal stress o 3 is almost vertical and its value corresponds to overburden weight (a z =12,1 MPa).

The first test considered was with fixed

Young' s modulus and Poisson' s ratio. In this case the results obtained with the two methods are exactly identical (table 1) to the exact values, even with a far initial solution. Thus, thèse methods can be used like the classical least square method v/hen the rock properties are well defined. Table 2 shows an example where ail the Table 1 parameters are unknown and the initial solution is close to the exact values. In this case, the algorithms converge to the exact values. However, v/hen the initial solution is relatively far f rom the exact values, the results are physically acceptable but the solution is not good. It shows clearly, that this algorithms must be used carefuly.

If we consider that the principal vertical

CONCLUSION AND FORTHCOMING RESEARCH

We tested thèse two algorithms with data coming from différent types of overcoring raeasurements. We didn't observe différences between results when we changed the norm. Nevertheless the évaluations number of F is less when condidering the euclidean norm.

We are allowed to work with the principal stresses and their orientations (expressed in term of Euler angles) rather than with direction and shear stresses.

The two methods reveal themselves to be very complementary and highly performant when the number of unknowns is less then or equal to six. In the opposite case, the results dépend mainly on the initial values, because of the high number of solutions [START_REF] Baucheron | Application de méthodes d'optimisation à la détermination des contraintesetdes propriétés géomécaniques d'une roche[END_REF].

The différent uses of the Powell's and Mifflin's algorithms showed that they fit very well with classic overcoring analysis methods such as least squares method. We now try to ameliorate thèse methods for future work. At first, the introduction of constraints on the unknowns (such as a variation interval) will allow to better understand the influence of initial values and to direct the parameters research towards given directions. The présent approach didn't give any informations about the précision of the results. So we are thinking of a probabilistical approach which will be based on the maximum likelihood : it means that we will work on the probability law of the unknow parameters rather than on their values. Finally, because we want to détermine ail the solutions of the équation F (E, a)=0, we use a process which combines the Powell's and Mifflin's algorithms with a Monte-Carlo method in order to see more precisely the solutions domains.

[7] POWELL MJ.D., 1964 -An efficient method for finding the minimum of a function of several variables without calculating derivatives. The computer journal 7 pp. 155-162.

[8] TARANTOLA A., 1987 -Inverse problem theory : Methods for data fitting and model parameter estimation. Elsevier Science Publishers.

[9] THEODOR R. , LASCAUX P. , 1986 -Analyse numérique matricielle appliquée à l'art de l'ingénieur. Tome 1. Ed. Masson, Paris. Détermine the m index which gives the maximum À.

Step3: Iff 2 >f 0 or 2(f 0 -2f t + f 2 )(f 0 -f t -A) 2 > A(f 0 -î 2 f then do not change the directions d v .., d a , set x° = x n and go to step 5.

Step4:li f 2< f 0 and 2(f 0 -2f t + f 2 )(f 0 -f t -A) 2 < A(f 0 -f 2 f then détermine the minimum of F in the direction x n -x°. The obtained point will be taken as new starting point x° at the next itération.

Replace the n indépendant directions in this order

Step 5 : If the stopping test is not verify go to step 1.

APPENDIX A

We describe hère the différent steps of the Powell's algorithm.

Step 0 : Initialisation : choose a starting point x° and n linearly indépendant directions d 15 d 2 ,.., d B , for example the canonic directions.

Step 1 : Détermine the séries of points x 1 , x 2 ,...,x n as :

F (x 1 ) = min F(x' +X d) = F(x M +X i d. ).
This could be done by one variable minimization like a dichotomy or a golden section method.

Step 2 : Define f 0 = F(x°), f 1= F(x n ) and f 2 =

APPENDIX B

The Mifflin algorithm require positive real numbers oc,p\Y,8 and p with r<l and (2P 2 n 2 y)<p. Given thèse parameters, the detailed algorithm is as follows :

Step 0 : Choose a starting solution point x of R-and a starting step-size s > 0. Set the séquence index k = 1 and the séquence values x 1 = x and s,= s. Tlien redefme x by : F(x ) = min [ F(x + td)

Step 6 : If F(x m ) -F(x) > -(ags) 2 , go to step 7. If F(x m ) -F(x) < -p^ll AF H 2 , choose some reduced stepsize r in the interval (0,s] and go to step 8. Otherwise set r = s and go to step 9.

Step 7 : Set r = s / 2 and x m = x.

Step 8: If x * x k replace k by k+1. Set the séquence values x* = x and s k = s.

Step 9 : Replace x by x m and s by r and go to Otherwise (as>i|AF|| andDqq-Eq<0) compute z as above and set d = -sign(z'AF)z.

Fig. 1 :

 1 Fig. 1 : Steps of overcoring method

  parameters are taken as data of the problem. The disadvantages of this method are that significant variations of the modulus values can occur from point to point in the measurement area, and that the properties of the rock can change between the moments when the in situ measurements are done and the détermination of the elastic constants (altération of the rock, humidity,...). The relation between the stress tensor (a) and the observed measurements (Y obi ) is linear. We hâve : = A.a where A is the compliance matrix.

  constants and stresses are separate variables this means that the values of the elastic parameters are not depending on the rock stress state. This hypothesis is very important and, in gênerai, implicitwhen the elastic constants are determined The idea is to consider E as unknown and détermine a and E so that Y^ = A (E).a the calculated displacements or s trains are équivalent to Y^. This problem can be solved by minimization of the function f (E,o) = ||Y^ YJI p where || designate the norm of the vector Y ob ^ Y^,.

[ 10 ]

 10 VAN MEERDEN X., 1983 -Stress Strain Relations applicable to overcoring techniques in transversaly isotropic rocks. International Journal of Rock Mechanics in Mining Sciences and Gteomechanics Abstracts. Vol. 20 -n° 6 pp. 277-282. F(2x n -x°). Calculate A = max[F(x.)-F(x, 1 )].

Step 2 :Step 3 :

 23 Stepi:Computeann-vectorof approximation fïrst partial derivatives AF by : AF. = (l/2s) [ F(x + se) -F(x -se.)] for i = 1 to n and an approximate gradient norm : || AF j| = Set the descent direction indicators o. for i=l to n : a. = +1 if AF<0 and O. = -1 if AF>0.Define a best axis point x by: F(x a ) = minF(x Compute an (n,n) matrix of approximate second partial derivatives A 2 F by :A 2 F U = ( 1/s 2 ) [ F(x + se.) + F(x-se ; ) -2F(x) fori= 1,2,.. ,n A 2 F y = ( a.o/s 2 ) [F(x + s^e. + sa.e.) + F(x) -F(x + sae.) -F(x+ saep ] for l<i<j<n Define a best corner point x c by : F(x c ) = min and a (possible) move pointx m by : F(x m )=min [F(x a ),F(x c )]. For 1 <j < i < n , if | A 2 F. | >y, replace A ;î F.. by "ysign(A 2 F..). Using a modified Cholesky factorisation procédure ( see [9] )rwicn compute the matrices L,D and E such that LDL'=A 2 F + E. Define index q by : D -E =min[D -E.]. If as > H AF || and D qq -E qq > 0, go to step 7 . If as < H AF ||, compute y 1 satisfying LDL'y 1 =-AF and set p = 1 ; and if E*0,sety 2 = -[il y 1 11/11 AF ||]AF and p = 2 and if D qq -E qq < 0 , compute z satisfying L"z = e and set y 3 =sign(z'AF)[}| y 1 1|/|| z ||]z and p = 3 , and define a search direction vector d by : d'AF + l/2d'A 2 Fd=min [ (y') l AF+ j ] withl<i<p.

Step 5 :

 5 Compute, if possible, a search point x + td, where t is a positive number satisfying F(x + td) -F(x) < pt [ d'AF + l/2td'A 2 Fd ] .

Table 3

 3 stress is known (x.x. o known), wich is a

	yz xz	z
	common assumption in many cases, the results
	obtained for the unknown parameters are as
	expected (table 4); the same if v is fixed anda z is
	unknown (table 5).

Table 4

 4

	parameter	exact values	initial values	MIFFLIN	POWELL
	E	75000	60000	75003	74999
	Ox	18,8	15	18,8	18,8
	Oy	23,1	15	23,1	23,1
	Oz	12,1	15	12,1	12,1
	*xy	2,1	5	2,1	2,1
	nb évaluations otF			162	653

Table 2 Table 5

 25


-in some cases, administrative authorities may ask a third party for a critical survey (etude de sürete ou analyse critique de l'etude de dangers) of the hazards and related intemal Operation programme.

In addition, general survey of an industrial estate is sometimes required by the administrative authorities.

In the first part of this paper, acceptable pressure and heat radiation threshold values and energy threshold values for missile effect have been choosen for the scenarios of major technological hazards such äs fires and explosions. The effects on the population and on the facilities are critically analysed and emphasis on energy threshold values for fragment dispersion will be given when presenting means of assessing the effects.

The four following parts will be devoted respectively to caiculation means used for evaluation of the effects of pool fires, bleve and explosion (both confined and unconfined) and toxicity effects in case of release of gas and vapors to the atmosphere.

Observations drawn from the investigations of accidents will be critically reviewed (8).

A more detailed review of several of these subjects is given in publication (6).

II -PRESSURE AND HEAT RADIATION THRESHOLD VALUES AND ENERGY THRESHOLD VALUES FOR MISSILE EFFECT

11.1. Pressure effects.

II.1.1. Effects on individuais.

After a dose examination of existing data, a threshold pressure value of 50 mbar was accepted for reversible pressure effects on man.

This level is consistent with the value fixed in french regulation related to safety distances in pyrotechnic plants (3) when using the pressure versus the scaled distance curves determined with trinitrotoluene detonations.

This french regulation takes also into account the threshold pressure value of 170 mbar for significant lethality.

These thresholds values are defined in terms of incident pressure. We do know this is an over-estimation for small explosions (for instance in the order of 100 kg TNT equivalent).

In fact, time has to be considered : the impulse effects have an important influence on men and facilities. For men, two types of effects are in fact to be examined :

-injuries to eardrums and lungs, -people set in motion by the shock wave.

It is well known that eardrums do not withstand a rapid rise in pressure. Time is a parameter to be taken into account by the means of impulse. Transient effects may appear when surpression exceeds 0.15 bar, if positive impulse -I -is over 0.02 bar.ms. Eardrums may be ruptured with a surpression up to 0.35 bar and an impulse over 0.5 bar.ms.

A sudden overpressure on the thorax may easily cause pulmonary injuries. Importance of the disease is an increasing function of the ratio P/Paim (overpressure/atmospheric pressure) and of the ratio I/M 173 / Patm^2 ^ere M is the weight of individual.

But people may also be set in motion by the shock wave. The initial speed is function of the overpressure P and of I/M 2 / 3 . When this speed is less than 3 m/s, there is no particular risk. By collision of the head against a part of a fixed installation, a fracture of the skull may happen when this speed is 4 m/s ; it will happen quite surely when the speed is 7 m/s and with 50 % probability when speed is 5.5 m/s. By collision of an other part of the body than the head, the corresponding values for speed are 6.4, 42 and 16.5 m/s.

Effects on structures.

Data are available about the behaviour of walls, structural equipment and window panes [START_REF]Comportement des vitrages aux effets d'une onde de choc[END_REF]. Structural equipment design may include layout and devices to limit explosion damage. Nevertheless the analyst has to bear in mind the synergetic effects a blast wave may produce on the surrounding equipments. Only few informations is available for instance in an existing plant, the main difficulty lying in the lack of data for previous designs.

With regard to the pressure effect on facilities, TNT curves (4) give the incident peak overpressures versus scaled distances. But the effects of reflections are to be considered ; these curves are consistent for explosives and mainly in the far fields. For detonation of gases [START_REF]Estimation des consequences d'un rejet accidentel dans l'atmosphere d'un produit toxique ou inflammable[END_REF], evaluation tests are summarized on figure l.

Currently, only scattered results are known [START_REF]Estimation des consequences d'un rejet accidentel dans l'atmosphere d'un produit toxique ou inflammable[END_REF] These threshold values are to be used only when men are unable to escape and run away from heat effects.

Nevertheless, when people is properly drilled, higher values could be accepted (for instance : 8 kW/m 2 for firemen with suitable equipment and a quite short stay).

For long duration effects, it is possible to quote correlations giving the threshold values of injury (at a low probability) and lethality äs a function of time (9).

In the case of BLEVE, the duration of the phenomenon is very short : it can be currently assessed that the duration is in the order of 20-30 s for spheres containing 500 m 3 to 1500 m 3 of liquefied gas. So, higher threshold values for heat radiation than those for high duration fires could be used.

Effects on facilities.

With regard to the heat radiation effects on facilities, various cases are to be taken into account. Thus, threshold values cannot be easily and univocally given. Many caiculations are possible when the behaviour of the material in presence of fire is known.

These considerations are global and, for example LANNOY (10) deals with heat conduction in concrete. But spacing distances derived from thermal radiation modeis are also needed. They are for instance given by CROCKER and NAPIER [START_REF]Thermal radiation hazards of liquid pool fires and tank fires" -I[END_REF] in table l. 11.3. Explosion energy threshold values for missile effect.

Energy threshold values.

For individuais, the only important energy threshold values for missiles effect is the one corresponding to lethality. It is not worth worrying about defining a reversibility threshold for when a missile penetrates any part of the body. The french pyrotechnic regulation gives this lethality threshold value äs equal to 20 Joules. This value is connected to the boundary between a zone defined by "serious and maybe lethal injuries" and a zone defined by "injuries".

FUEL TANK DIAMETER (m) *S/D RANGE

When an explosion occurs, the impulse Jp.dt which sets the fragments in motion is of paramount influence.

In industry, most accidental explosions experience proved that the mean weight of a fragment is of the Order of 30 to 100 kg. When the fragment has a 10-15 kg weight and a 4 m/s speed, its kinetic energy is 80-120 Joules. Thus we have choosen an energy threshold value equal to 100 Joules.

Missile effects.

For the prediction of the missile effet during accidental explosions, simple caiculation methods were defined in TNO yellow book (18). They were completed in the UCSIP (Union des Chambres Syndicales de l'Industrie du Petrole) guide for hazard evaluation in the petroleum industry (17).

This approach can be summarized in the three following Steps :

-energetic caiculation of the initial speed of a fragment, -ballistic caiculation with possible consideration of a drag coefficient, -comparison of the ballistic results to lethal energy threshold values for individuais and to impact Perforation speeds for pieces of equipement.

Even though complex softwares have been developped, the mechanisms of bursting for different types of material and vessel shapes was only validated until now for simple configurations and mainly metallic materials under dynamic loading from an explosion.

Therefore the first main difficulty is to define the bursting pressure of the enclosure, the location where the initial breaking occurs and the number and shapes of the missiles.

Then during the projection, the direction is strongly dependent on pressure history and the shape and mass of the missiles.

But, for apparatus of complex shapes and buildings, directional effects and the exact location of the weakest part of the System combined with the propagation mechanism of the explosion imply great uncertainties about the maximum landing distance for the missile and its random distribution.

Therefore, comparisons of results gained by investigating accidental explosions and by simple caiculations are needed for prevision purposes in hazard studies (see V).

III. POOL FIRES.

The main purpose is to assess the radius of the effect (r). Buming rate, flame height, amount of heat radiated from the fire are important parameters.

For LPG (Liquefied Petroleum Gases), we have considered an emissive power of 60 kW/m 2 according to MIZNER and EYRE (16) (to be compared to 30 kW/m 2 for liquid hydrocarbons).

From this thermal radiated flux one can caiculate the flux received considering view factor and attenuation through air can be caiculated. The table 2 gives caiculations of the radii for different heat radiation threshold values with 20 and 30 m long square pools.

These caiculations are conservative and other parameters such äs the geometry of the dyke, screen effect of construction, view factors are to be taken into account if more accuracy is needed.

The influence of the wind could be also of paramount importance leading to the tilt and the drag of the flame.

Then, the radiative effect may be caiculated from a tilted cylinder in the wind direction.

The area involved takes an elliptical shape. Even, the longest dimension is only 25 % higher than the previously caiculated radius. But, the convective effect may become predominant.

Substance

IV -BLEVE.

This phenomenon, by its thermal effects, can trigger the most important major technological hazards. Comparisons between thermal and pressure effects are given.

For example, for propane or butane spheres (the volume of which ränge from 500 to 1500 m^) the following caiculations can be made.

IV. l. Thermal effects of Bleve.

Differents means for assessing the duration or the diameter of the fire ball and its radiation have been developped by different authors such äs : NAZARIO (14), UCSIP (17), BAKER and al. (4), TNO (18

)
For example, TNO (18) proposes to caiculate the duration of the fire ball and its radius using the formula : duration (seconds) t 0.852 M 0 -26 radius (m)

3.24 M 0 -325 M (kg) is the content of the vessel Assuming the heat radiation on the surface of the ball is 200 kW/m 2 , and using the threshold curves reported by K.S MUDAN (9) about fatality levels for thermal radiation, it is possible to caiculate : radius (m) for l % lethality 3.12 M 0 -425 radius (m) for significant bums 4.71 M 0 -407

These radii are caiculated at ground level, without attenuation by air.

For a 1500 m 3 butane sphere, the radius for l % lethality is 982 m and 1162 m for significant bums.

We have considered the existing empirical correlations giving the radius and the duration of the fire ball and its radiation and compared these caiculations.

When comparing these results, we föund large discrepancies (mainly related to the values ofemissive powers, the position ofthe fireball, the attenuation by air).

IV.2. Pressure effects ofthe Bleve.

According to TNO Yellow Book (18) the peak overpressure of a shock wave with adiabatic flash for hydrocarbons is related to the scaled distance r/(2M)°-33 with the parameter "superheating" (liquid temperature of the content of the vessel just before bursting open -the atmospheric boiling point of the liquid).

For example, assuming an overheating of 150 K, it is possible to caiculate the radii for pressure threshold of 50 mbar.

In this case, the radius (m) for reversible wounds (pressure threshold value = 50 mbar) = 8.70 (m) 1 / 3 .

For a 1500 irP butane sphere, the radius for significant bums, given in the above mentioned table, is 1162 m while the radius for reversible injuries (pressure threshold value 50 mbar), caiculated by the above formula is 792 m.

It can be concluded that for injuries the pressure effects radii are lower than those for heat radiation. Consequently, for the protection of the population, (if we do not consider the missile effects), one need to consider mainly the radiation effects. But it might be pointed out that, until now, the delay of occurence of a BLEVE is quite unpredictable.

V -CONFINED AND UNCONFINED EXPLOSIONS.

Accidental explosions may be either confined or unconfined with hazardous substances äs diverse äs explosives, propellants, fertilizers, combustible gases and vapors and inflammable dusts.

With gases and vapors, DA VENPORT (12) pointed out that only a small part of the released gas is generally involved in the unconfined explosion.

Such an explosion occured some years ago in France in a chemical plant after the accidental release from a vessel under 40 bar pressure : 160 kg of a mixture of H^-CFŵ ere released, after the breakage of a connecting duct. In this particular case, from the effect given by the blast wave and the mechanical damages it may be concluded that only l kg of the H^-CI-H mixture released was involved in the explosion. After dose examination of the mechanical damages, the TNT equivalent could be estimated at 5 kg.

INERIS established a code describing the Variation with time of the dimensions of the inflammable cloud and validated it by testing the horizontal discharge in air of a vessel pressurized with H^ or 0-4 (13).

In the survey of the hazards for assuming the maximum effects of the explosion, the time lag at which the inititation of the explosion occurred is äs important äs the source location in Order to define the safety distances. Unfortunately, few codes can describe the generation of a cloud, particularly when it results from the discharge in air of a vessel pressurized by an inflammable gas. Some recent accidents in France in a gasoline storage plant (St Herblain 1991), in a refinery (La Mede, nov. 1992) and in a LPG storage plant in USA (Brenham 1992) emphasized the need to develop more convenient modeis to deal with such unconfined explosions.

The European programme MERGE (Modelling Experimental Research into Gas Explosion) deals with modelisation and experimental validation on unconfined explosions of inflammables atmospheres.

The explosion can be semi-confined, äs in 1986 in a casting line, where a very violent aluminium explosion occured causing the death of 4 people, injuring 25 others and with extensive material damage. This explosion triggered by a thunderbolt involved the Vaporisation and atomisation process of liquid aluminum. The resulting vapors and droplets, mixed with air ignited by the thunderbolt gave an explosion of and above the casting line (ground level). Then, a second explosion, not äs powerfui äs the first one was initiated in the casting pit by the action of liquid aluminium on water. By examination of the damages to the casting pit and to the surrounding buildings, evidence was gained that the TNT equivalent was about 200 kg and 100 kg respectively for the first and second explosions. Pieces of equipment of the casting line were thrown up to 700 m (15). Every year, a lot of accidental confined dust explosions occur in elevator buildings and silos. Damages can be more or less extensive but in some cases, when the confinement is important, with rather resistant concrete buildings, concrete missiles are thrown. In 1982 a barley and malt dust explosion occurred in Metz, France and induced the collapse of 9 out of 14 concrete cells (äs a consequence of the fact that there was quite no links between the cells -independent cells-). A violent explosion in the work tower resulted in a shock wave with the projection of lightweight panels blown in a verydirectional area (no more than 100 m). The dispersion of very important concrete blocks was not larger than 80 m corresponding to the maximal height of the elevator.

The same year, another explosion was initiated in the upper part of a sugar storage plant (2 x 20000 t cells and l x 40000 t cell).

The two 20000 t cells were covered with an expanded concrete roof and during the explosion, quite small pieces were thrown up to a distance of 500 m, but the vertical walls remained Standing even though some crevices appeared. For the largest cell (40000 t) the reinforced concrete roof first was raised up with some dismantling in large pieces but then feil again into the silo cell.

With these examples of explosions, it became evident that there is a lack of validated methods to predict the effects of such accidents. Efforts must be devoted to validating the existing modeis or to improving them.

VI -TOXICITY EFFECTS IN CASE OR RELEASE OF GASES OR VAPORS TO

THE ATMOSPHERE.

An accident on a pipe or on a tank containing compressed or liquefied gas, a fire with chemicals may result in releases of gases or vapors mists or soots and dusts to the atmosphere. In a fire the decomposition speed of the products and the nature of toxic emissions issued must be defined (7), but will not deal with the topic of the present paper.

To evaluate the consequences of a release, in terms of effects zones, parameters about the release (source term, dispersion conditions) and about the effects on human being have to be identified.

With regard to the source term, from tanks or pipes, the release may be continuous (permanent if flow rate is not limited in time) or instantaneous depending on the type of rupture : for example fissuration in a tank or total rupture of this tank. The source term is also influenced by the physical state ofthe content before release.

The dispersion conditions of the toxic cloud depend on the atmospheric conditions (stability, temperature, windspeed ...) and on the site conditions (roughness ...). The phenomenon has to be studied under unfavourable atmospheric conditions.

For effects on human being, when some people inhales an atmosphere polluted by toxic components, the effects usually studied are the occurence of lethality (with low probability), the occurence offaint, ofcough ...

For people exposed to such atmosphere, these effects are function of concentration (C) and time of exposure (t).

For a constant effect E, a relation ship between concentration and time can nearly be found : E = C" x t.

Where C is the concentration Step t is the time of application of C n is a constant depending on the component The effect E will occur if E s C" x t. Thus, it is important to be able to caiculate concentration field issued from the release and also to determine the history of concentration at each given point downwind the point of emission.

In case of a permanent release, the zone to be considered for a given effect is the same äs the zone where the concentration is equal or above a given concentration. For this type of release, this is due to the fact that concentration versus time begins to increase and then remains constant when the System is established.

In case of an instantaneous release, concentration versus time is not constant, and the concept of equivalent dose defined above has to be considered.

Probit equations giving values for C, n, t are proposed to be used when the concept of equivalent dose is retained. The approach is very attractive but, depending on the n value choice, important deviations can be introduced in the distances caiculated. On an other point of view, for the effects on human being, care has to be taken for the validity of the data, especially in case of correlations with results from experiments on animals.

At last, number of dispersion modeis have been computerized, but all of them have limitations and need care for their use in given applications.

VII -CONCLUSIONS

The critical analysis of hazards studies and accidents connected to pool fires, BLEVE and confined and unconfined explosions pointed out many lacks of information to evaluate the pressure, heat and missile effects.

The acceptable pressure and heat radiation threshold values and the energy threshold values for missile effects are generally based on experience gained after accident investigation or intentional experimental explosions.

The effects on individuais and facilities are both to be considered separately.

For pressure threshold values, the effects on human beings are quite well understood ; but for facilities, only scattered data are available.

Heat radiation threshold values are to be choosen differently whether BLEVE or large pool fires are involved.

For the maximum landing distances of missiles energetic and simple ballistic caiculation can give an order of magnitude for bursting of vessels. In other cases only results gained from accident investigations are useful.

Caiculation means for pool fires are consistent to evaluate the heat radiated from LPG (Liquefied Petroleum Gases) and liquid hydrocarbons fires. The effect of convection needs more consideration.

For BLEVE, current caiculations are sufficient for predictive purposes but the delay of occurrence is rather unpredictable.

For toxicity effects, number of computerized caiculation modeis can be found, but their use often need to be supervised by a specialist.

More experimental investigations for both confined and unconfined explosions are needed to get a better understanding of the mechanisms of the explosion, and of the behavior of equipment and buildings under dynamic loads given by these explosions.

Figure l :

 l Figure l : Detonation peak overpressure of unconfined air-hydrocarbon mixtures (spherical geometry) [from A. Lannoy (5)]

  For l % fatality level, the threshold flux : (| ) = 190.811-0 -771 ((() kW/m 2 and t s) is given on figure2according to EISENBERG values quoted by MUDAN (9) .kW/mF igure 2 : Fatality levels for thermal radiation (Eisenberg and al.43) [From MUDAN (9)]

  Radius (r in meters) from the edge of the pool at different heat radiation threshold values Nevertheless, the tilting angle and the flame length are usefui to determine the pieces of equipment engulfed in fires.


Véronique Soukatchoff

Rafik Hadadou

Didier Driancourt

Yves Paquette

Observation and auscultation of the geotechnical behaviour of a slope in an open cast mine influenced by old underground mining (South-western part of France)

Y.Paquette

INßRIS, Verneuil-en-Halatte, France ABSTRACT: The final slopes of a French open pit mine of the "Charbonnage de France" Company which final depth will be 300 metres, have been designed after a Standard geotechnical survey. This one has taken into account the geomechanical, hydrogeological, structural parameters äs well äs the" decohesion", induced by former mining subsidence. However some slopes can locally present risks of slipping (increasing with the depth of the pit) induced by old Underground mining. A methodology has been defined conceming the slope auscultation and its behaviour including the use of numerous measuring tools (topographic measurement, inclinometer, equipped cable bolts...). Two auscultation cases are presented, one on a slope slipping, another one on a potential slide. In the first case, the instrumentation checked that the reinforcement is sadsfactory. m the second case, the instrumentation allows to analyse that the correct measures have been taken to prevent any risk.

l -INTRODUCTION

The Sainte Marie open pit is a coal mine of the Mining Unit ofTam, H.B.C.M. (Houilleres du Bassin Centre Midi), in the south-westem part of France, near Albi (Fig. l). In this area a large amount ofcoal has been exploited by Underground mining. This pit was designed in order to exploit the coal remaining around the shaft (Saint Marie shaft) ofan old Underground mine situated in the basin of Carmaux. The diameter at the top of the pit is 1200 metres and its final depth will be 300 metres. The first 100 metres are composed of tertiary deposits (clayey sand) which covered the carboniferous formation. The average slope angle ofthe Tertiary is 37° (withoutbenches) and in the Coal Measures it was foreseen from 37° to 50°( with benches of6 metres high) depending on the slope situaüon. At present üme, the depth of the mine is about 160 metres. Nine coal seams (named form the top: C, D, DT, F, E, G, H, I, J) have been mined by Underground working between 1900 and 1984. Different methods have been used depending on the thickness, the dip of the layer and the dimension of the panel. In fact panels were backfilled, caved or undermined long-wall. The basin of Carmaux is a large synclinal split by a dense network of faults which directions are approximately N 140 E. The dips and the dip directions Fig. l : Situation of these principal normal faults are quite variable because of a complicated succession of tectonic phases. The throw ofthe faults oriented N 140 can be from 20 to 100 metres, and the spacing is from 10 to 100 metres. This particular network ofmain faults leads to consider the basin äs a succession of narrow compartments. The structure äs a large influence on the slopes behaviour. In fact the centre ofthe open pitis composed of unexploited layers of coal, because it is the coal which was left around the shaft, but, dose to the slopes, begin the old exploited long walls. These long walls are at different topographic levels due to the particular structure and have been exploited in panels lined by the faults odented approximately N140 (Fig. 2). The first design of the open pit was done by a Standard geotechnical surv ey; this one has taken into account the geomechanical, hydrogeological, structural Parameters äs well äs the "decohesion", induced by the revival ofsubsidence due to old Underground mining. However some mining slopes can locally present risks of slipping induced by old Underground mining. We shall present, at first, the different kinds of instruments installed in the mine, then we shall focus on two cases: one on a slope slipping, another one on a potenrial slide.

TYPESOP INSTRUMENTATION

As far äs a lot of the equipment is well known, we shall briefly describe the principal types of instnunentadon and only insist on the ones which may be more unusual.

Topography

About 60 topographic points are installed all around the mine. These topographic plugs are more nmnerous in the areas where slope stability problems were detected. When there is no particular problein the Position ofthe points are monthly collected. In case of movement the monitoring is more frequent and it can lead to make two measures a day when a sliding has been detected (cf. §3). The topographic survey is done with the help of a T 2000 theodoUle «Distomat Wild». The plugs are fit out with prismatic reflector. The accuracy is 1/10000 grades for the angles and 5 mm added to 5 mm/Km for the distances.

Inclinometers

Ten vertical inclinometers with depth going from 35 to 70 meters have been installed in the mine. The inclinometers are localised were slope problems were detected. We shall present the measures of this equipment with the description of instabilities.

Extensometers

Somerodextensometers(Wr-Flex)are setin boreholes. In general, the equipment we use, has 5 anchor points, 5 metres away from one another, so the total length infonned is 20 metres. The first anchoring point is generally at about 20 metres from the surface. The extensometers are vertical or tilted and are manually measured.

Shear strip

The shear Strips which are used and that we call «brealdng tube» (made by INERIS) are composed of electrical resistances regularly positioned (in our case each 50 centimetres) and linked with two copper wäre wires. The all package (electrical resistances and wires) is included in a PVC tube filled up with glass spheres and resin. An element of shear strip has 12 metres long and it is possible to put several elements linked to each other in a hole. Two measures can be made: one from the top and the other one from the end of the tube. In case of movement, the shear strip is broken and the measure of the resistance of the tube falls down. As electrical resistances are installed each 50 centimetres, we can localised a rupture with this accuracy.

Tension-measuring gauges

When cable bolts are installed in Order to reinforce the slope, some Tensmeg-70 gauges (Supportek Inc.) are used. The working principle of the gauges is based on the change of the resistance of the resistance wire caused by cable elongation. Calibration tests have shown that there is a linear relationshipbetween readout values and tension applied to the cable. The value of resistance aUows to know the tension in nie cables. tn July 1990, after atopographic reading, the beginning of a movement was detected at the interface between Tertiary and Stephanian (by the displacement of the spot 500) in the east part of the open pit. At this place the pit was 150 metres deep and 4 benches of 6 metres high, of Coal Measures were mined. Opened cracks were detected in the haulage road (Fig. 3) above the plug 500. The analysis of this movement and the information from the Underground mining, show that the sliding mass was just above an exploited panel of the E seam (Fig. 2) and was delimited, in the coal bearing ground, by two faults : in the north, a fault oriented N 150 E, W 55, named n° 40, and in the south a fault onented N 125 E, N 44, named n° 60 (the number of the fault was assigned during the Underground mining). The panel ofthe E seam was a long wall exploited in 1960 and backfilled with shales without any cohesion.

The movement was finally stopped by constructing a embankment-buttress, 35 metres wide and 40 metres high.

In order to monitor this sliding and verify the efficiency of the buttress, the following apparatus were installed (just after the detection of the movement):

-different topographic points are in this part of the slope: n° 417,500,507, 515,522... (Fig. 3); -3 shear Strips in different holes, 45° tilted. They were respectively 10 to 22,20 to 32 and 27 to 39 metres from the top.

-2 cable bolts of 24 metres long (250 kN of ultimate tension load) equipped of5 Tensmeg gauges. The first cable from the elevatfon 154 was 30° tilted, the second one, installed form the elevation 182, was 25° tilted (Fig. 4 As it took ten months to build up the embankmentbuttress to its necessary size, the movement was

mm oa (A ib i 12^001^1 IK d M 'l f Shw i 3 (30 m) firips j----------! 2 (24.5 H)) Ehe die ..-,, rupmreat Nc ä ------------I.ämeüW I^C5 ------~-----)-------- , •l | S 3 s l Is J l l • • /i i:;. :: ',' I 2 3 5 [NC8 -----• niptuneit -22n>e[rcs ^ 5 ------^. 4 2 ' i i i 06 07fll08»l
Fig. 7 : Time evolution of the monitoring gradually stopped. The displacement of the spot 500 (Fig. 5) shows the time evolution of the movement. We can see that after the embanikment-buttress was finished (May 1991), the displacement was less than a few cenümetres. Most of the apparatus has shown an important acüvity ofthe slide between July 1990 and May 1991. The phenomena is shown by failure of shear Strips, deformation of the inclinometers, values ofthe gauges (Fig. 6) on cable holt which indicated that the load have exceeded the ultimate strength of the cable : 250 kN (in general the failure of the cable bolt airive for 9000 micro strains). The time evolution of the phenomena is summarised in the Fig. 7. After May 1991, wecan control that the movement was stopped, because the evolution ofthe topographic points is very weak and the mclinometer INC8, instaüed in August 1991 was not deformed.

The monitoring shows that the slickenside was effecüvely composed by the E panel and by a curved surface in the Tertiary formation,just under the haulage road. Different levels of rupture were detecied. They showed that the sliding mass was deformable, and that the movement cannot be modelled äs an undefonnable volume. A modelling has been made with the ÜDEC code, it showed that low cohesion of the backfill of the E seam and the network of fault can explain the extend of the slide.

By now the area is still monitored; no important displacement has been detected since 1991.

AUSCULTATIONOFARISK

In the north of the slide, we described previously, there was another configuraüon quite similar: between the two faults N° 40 (north Umit of the previous phenomena) andn°100(N117E,NE48),apaneloftheGseamhas been exploited in 1960 (Fig. 2). The dip direction ofthe exploited seam is, äs in the previous case, oriented to the centre of the pit (the dip of the panel has the same direction äs the slope). In order to make sure this configuration will not induce a slide, monitoring was installed (summarise in the table l) The auscuitaüon ofthe G seam showed that there was a movement which progressed since May 1991. This movement has been detected by the topographic point 515 (Fig. 10), the extensometer N° l (Fig. 8) and the inclinometer INC7 which showed significant levels of deformation at 28 and 34 metres deep, in May 1991 and was finally cut in May 1992. Rod extensometers were very usefui to detect movements underthe slopes, increasing with the depth of the pit. Some reading are someüme difficult to explain because they are the result of different movement: revival of subsidence, settlement due to the embankment, shear movement on the old worked seam. As the extensometer, the inclinometer and topographic points showed an acceleration of the deformation in this area, reinforcement was started up at the beginning of1992. In this case, again, in order to model the beginning of the movement and to caiculate the reinforcement, we needed to take into account the result of the different kinds of monitoring, to have a better understanding of the movement. The results of this analysis lead to reduce the slope angle.

CONCLUSION

We analysed two cases of auscultation. In the first example, a sliding was detected by topographic measurements. In order to analyse the deformation and to understand the development of the movement, different kinds of apparatus were installed. In the secondcaseariskof sliding was detected byageometric analysis and it was monitored. Different kinds of monitoring System were used. The topographic measures represent a good alarm, which can detect a movement and allow to define the limit of a slide at the surface, but they cannot explain the mechanism ofa slipping and the volume involved. The shear Strips allow to detect a rupture, but they work only one time. The borehole extensometers show displacements, but the results of the measures are sometime difficult to explain when we have interference of subsidence, settlement and shear displacement. The inclinometer is the only apparatus which can indicate the direction ofthe movement. The tension-measuring gauges allow to verify the cable bolts are acting. In order to have a correct understanding of a whole movement and to model it, we need to associate the different kinds of monitoring System. By now, two other cases of potential problein have been detected. These configuraüon are a little different from the others because two superposed exploited seam will be cut by the slope. These cases will be monitored with the same Idnd of instrumentation.

  Observation and auscultation ofthe geotechnical behaviour ofa slope in an open cast mine influenced by old Underground mining (South-westem part of France) VSoukatchoff Laboratoire de Mecanique des Terrains, Ecole des Mines de Nancy, France R. Hadadou INERIS, Laboratoire de Mecanique des Terrains, EMN Nancy, France D. Driancourt HBCM, U.E. TARN, France

  Fig. 2 : Schematic map of the open pit

  ) -3 inclinometers : INC5, installed from the elevation 226 and 60 metres long, INC8 installed from the elevation 226 and 38 metres long (this apparatus replaced the INC5 after this one has been cut), INC6 installed from the elevation 182 and 40 metres long.

Fig. 4 :

 4 Fig. 4: Section of the east slope through the exploited E seam

Fig. 5 :

 5 Fig. 5 : Time displacement (displacement in metres) of the poim 500

Fig. 8 :Fig. 9 :

 89 Fig. 8 : Time evoluüon (displacement in mm) of extensometer N° l


R Bigarre

K Ben Slimane

J Tinucci

3-Dimensional modelling offault-slip rockbursting

A research program has been carried out at INERIS aiming to quantify rockburst potential from mining-induced fault-slip. As a part of the research, numerical modeling of fractured rock mass has been undertaken, using the three-dimensional distinct element code 3DEC. Results presented in this paper demonstrate a very good agreement between caiculated defonnations of modeled faults and the experienced rockburst sequence of the Estaque-sud district of the colliery.

l. INTRODUCTION AND GENERAL SETTING

At the Provence colliery, coal is mined at a depth reaching 1100 m. The seam thickness is around 3 meters, while strata dips westward around 10°. The longwall face method with caving process is used, involving high mechanization and self-advancing support for faces of200 m ofspan. Rate ofproducüon has increased steadily all along the past, reaching now the value of 11 tonnes per shift and per day, with a average, daily rate advance of 6 meters per day and per working faces.

Nowadays, the coal mine experiences a daily average of 20 seismic cvents of magnitude 1.5 and grcater, 15% of which are magnitude 2 and more. Most of these events are attributed to the goaffing process associated with the longwall mining Operation. However, on an annual basis, many of these events result in serious rockbursting damages at the advancing face and along haulage gateways. As regards the southern part ofthe colliery and the mining of Estaque-sud district.which began in 1987, many major tectonic faults have been suspected to play a major part in dynamic loading of the coal seam through fault-slip induced by mining. flg. l-Geological cross-section ofthe basin

The general geological setting of the basin is quite simple. After Caviglio & al [1988], the structure is overridden by a major thrust sheet overthrusting northward, with a average dip of 25°( figure l). Strike-slip, sub-vertical faulting is present all over the area, with lengths ofseveral hundreds of meters. Two major zones have been distinguished in the coal field, with regard to the direction ofthe strata and stress measurements. In the zone we are concerned with, strong anisotropy ofthe principal stress components has been measured many times, characterized by high horizontal tectonic Stresses and a sublithostatic vertical stress. Both have been explained by the regional, geological history (Piguet & Georges [198l], [START_REF] Revalor | Seismic and seismo-acoustic experiments applied to the prediction ofrockbursts in French coal mines[END_REF], [START_REF] Gaviglio | La deformation cassante dans les calcaires fuveliens du bassin de l'Arc. Comportement des terrains et exploitaüon miniere[END_REF]). Associated strata are made of limestones, qualified äs hard and brittle [START_REF] Josien | Lutte contre les coups de couche[END_REF]).

ROCKBURST MECHANISMS AT THE ESTAQUE-SUD

DISTRICT

For the last fifteen years, stimulated by the steady increase in the daily mine tremors and annually rockburst occurences, a research program has been undertaken at INERIS, aiming first to understand the mechanisms involved and then to improveprevention. Classified with regard to rockbursts lucations and effects at the Provence coal mine, three main types ofbursts have been recognized, äs [START_REF] Gaviglio | Tectonic struture, strata properties and rockburst occurence in a french coal mine[END_REF]) (figure 2):

-type l: ends of the faces, especially on the old panel side. These bursts are now largely controled by means ofdestressing holes (figure 2a), although this method lacks of accuracy, -type 2: coal bumps, buckling ofthe floor, more current at the present time, over length sometimes greater than a hundred meters, can affect the gateways either ahead ofthe face (old panel side) or behind the working face, at a distance ranging from 50 to 150 meters (figure 2b) -type 3; strain bursts in unmined, overloaded stiff pillars (figure 2c)

Seimic energy associated with rockbursts varies around 10 8 -10 9 Joules, with an associated Richter magnitude ranging from 2.2 to 3. At the Estaque-sud district of the mine (figures 5-6), mining started in 1987 with longwall T13. During the 3 years following period, with a span ofone panel wide (200 m), then two (400 m) and three (600 m), 22 rockbursts 1 were recorded, starting with the mining of the second panel T14, most of the events being of type 2. A schematic description ofthe larger damages is suggested in figure 3, with following characteristics:

-violent expulsion of the coal in the gateway, -no significant fracturation or convergence of the immediate hangingwall, -quite often accompanied by floor heavage reaching l meter (whether due to buckling mechanism or deeper shear failure is still not clear [START_REF] Mathieu | Apport de l'ecoute sismoacoustique pour la surveillance des chantiers miniers affectes de coups de terrains[END_REF]). It is worthy noting that this kind ofdamage has been controlled for the last year by floor slotting ahead of the face, although the efficiency of this method has not been accurately estimated, due to the lack of data since its implementation.

l) We include here all significant dynamic events recorded, ranging from dynamic Spätling to large, Underground damages described here-after. The main hypothesis for type 2 events tends to classify them more precisely äs rockbursts triggered by dynamic loading generated by large mine tremors, induced by tectonic fault-slip or sudden failure of stiff bedding planes in the high roof. Naturally, Potential means ofconfinning this kind ofmechanism are very few because of the difficulties to get data. Extensions of the mining areas, poor access to faulted areas (one mined seam) and poor understanding of the roof behaviour do not permit to get valuable Information. Two types of investigations have therefore been undertaken:

-developing a mine-scaieseismic network, ableto give location ofeach mine tremors with good accuracy äs well äs its energy and seismic moment. This should permit to relate the located focus ofthe mine tremor and the Underground damaged areas and thus assert which mechanism of rupture may be considered. This has been undertaken two years ago (Ben Slimane & al [1990]). INERIS is currently isnproving the network to get accurate locations and better focus parameters, -analysing with all avalaible data the major, suspected faults respectively with mining geometry andscenarioto getabetter understanding of potential fault-slip behaviour. This has been undertaken recently and use of numerical methods is presented in this paper.

NUMERICAL MODELING

Numerical modeling has been carried out aiming to quantify rockburst potential for the seismic triggering mechanism from faultslip along major, pre-existing geologic structures. Because ofboth the mining configuration and orientation ofthe faults oftheEstaquesud area to be modeled, it was chosen to undertake three-dimensional numerical analysis. Eventualiy, the strongly discontinuous natureof the problem conducted us to the choice of the distinct element method.

Due to the lack of seismic data over the period of mining ofthe Estaque-sud district and the insitu conditions for mined areas of such extents, the aim of this study was to:

-examine the ability ofthe three dimensional distinct element method (3DEC, Itasca) to study fault-slip assessment for a complex System ofdiscrete, deformable biocs, -examine the fault-slip potential for large-scale fauits iying in the roined area and correlate in space the modeled mining process and the incremental plastic deformations with the insitu recorded rockburst sequence, -to bring forward a methodology of modeling cäoseiy associated with geological survey and above all with data froin the newly settled seismic network available.

3DEC SOFTWARE

3DEC is a PC-based Computer program using the distinct element method and a centra! finite difference scheme to simuiate the mechanical response of three-dimensional blocky Systems. Handling either rigid or deformable blocs, the formulation used permits to simuiate large displacements and rotations of the biocs relative to one another, including detection of new contacts, whiie the solution scheme is explicit in time. During each increment of time, Newton's law of motion is used to obtain velocities and displacements from the unbalanced forces. Mechanical caiculations may be described äs in figure 4, which shows the importance ofthe contact logic implemented. A complete description ofthis and oftbe caiculation cycie are given by [START_REF] Cundall | Formulation of a three-dimensional distinct element model -Part I a sheme to detect and represent contacts in a[END_REF] and Hart & al [1988] respectively. Note that when deformable blocs are used, modeied joints are sudivided in subcontacts corresponding to the finite difference tetrahedral zoning ofthe faces, while each surface node is the centroid ofan area defined äs the subcontact. This one keeps track of the interface forces äs well äs slipping or Separation. Graphical interface is largely developed, permitting to model äs efficiently äs possible well-conditionned problems compared to their original complexity.

Modeüng of rockburst mechanisms with 3DEC has been undertaken before, to simuiate fault-slip behaviour of discontinuous medium, applied to fault and dyke slip at the Strathcona mine, Canada (Hart & al, [1988]), [START_REF] Tinucci | Assessment of seismic fault-slip potenüal at the Strathcona mine[END_REF]). The numerical analyses were able to point out the consistency of fault-slip assessment in mine-induced seismicity and rockbursting. Initial pre-mining state of stress is chosen to be very dose to available field measurements obtained in the Etoile-sud district, dosest to the one modeled. Values and orientations are noted in table 2.

Two parameters are quantified in order to relate plastic deformation along each feature to each sequential excavation:

M = Z. A D
where As is the subcontact area and Du SubC.

its tangential displacement. M may be interpreted coarsely äs the seismic moment ofthe fault divided by its shear modulus. It characterizes the mechanical moment acting on the structure while new equilibrium is reached.

-AE =~^,FyD^ where Fg is the tangential force acting 2 SubC.

at the subcontact location. AE is the non-recoverable, released energy dissipated by the excess shear force induced at each step.

These two parameters are related, in such a multi-step, quasi-ÜJE static analysis by the relation: M = --where T is the T tangential stress acting at the subcontact location.

NUMERICAL PARAMETERS

Our 2800xl600x2400m model of the rock matrix consits of 188 convex blocks formed by 7 structural features, comprising 5 large-scale faults (500 to 2000 m of length, figures 5-6)) and two bedding planes (dipping east-west 10°) located at 150 m in the upper roof(representing coarsely the Fuvelian strata), and below the seam, and in thefootwall respectively. Four longwalls (T13, T14, T25 and T 15) are simulated, with a mining scenario reproducing the in-situ excavaüon process geometry. The four longwalls were then excavated in 15 incremental Steps, made of deletion of blocs, (figure 6) with equilibrium reached at each step, providing a quasi-static analysis able to put forward the influence ofthe incremental mined areas on the plastic deformations along the modeled faults. Each step represents an excavated volume equivalent to two months of mining at a rate of 100 meters per month.

Defonnable blocs, zoned by around 70.000 finite-difference zones, are assumed to behave elastically, while all structural features follow aperfect.'elastoplastic behaviour, based on aMohr-Coulomb yield condition (table l). Figures 7 and8 indicate the energy dissipated through plastic strain and the parameter M at each step both with the rockburst sequence plotted on the right, vertical axis versus the step number ofthe simulaüon. Results show that only the simulated overthrusting fault (fault A) and upper bedding plane (feature I) show large plastic deformations. These deformations appear above all frorn starting of longwall 14 (mining step 8-9). Summing up briefly, we can do the following comments :

-the qualitative correlation between M for both strutures and rockburst sequence shows a good agreement -rupture mechanism along fault A is due to shear failure, induced by both decrease in normal stress and increase in shear stress, coupled with stress tensor rotation. The amount fig. 6 -modeling ofthe Estaque sud district faults are indicated with dip direction and dip Geometrie projections of locations of maximum shear displacement (figure 9) on the seam plane are approx imately, verticaliy plumbed with the step excavation. Energy dissipated at step 9, if caiculated on a daily basis, provides a value of S.IO 9 !, which corresponds to an event with a order of magnitude of 3.

3-

failure mechanisms along bedding plane I are of two types: shear and tensile due to flexural behaviour. The relatively small energy dissipated is due to a low induced shear stress on the fault, parallel to the seam, and dipping 10° with regard to original principal sü-esses. In fact, tensile failure takes place essentiaily at steps 30,11 and 15 with Ehe widening of the mining area.

Slip along fault A is essentiaily oblique, i.e. with a mixed offset ofreverse and right-lateral strike faulti ng. This motion corresponds to the natural, thrust faulüng of the structure. Interaction of the two structures A and I dose to their intersection is difficult to estimate. It is worth noting tliat they most likely amplify each other because of the compatibility of the kinematics of the blocs.

Along other features, there is no noticeableplastic deformation although induced Stresses are unfavourablefor mostofthem, i.e. ratio T/o increases, except significant deformation along fault C. However, if large strain was to be obtained on this modeled joint, it would conduct to reduce friction to unreaiistic values (20-22°).

4-

SUMMARY AND CONCLUSIONS

In fact, the aim of this study is to evaluate whether using modeling oftypical, discontinuous problems in prediction of faultslip rockbursting might be useful. This study shows an good agreement between the rockburst sequence and the response of some of the discontinuous features lying in the mined areas. Mechanisms involved in the response of the System are clearly identified, critical geometries and Spans are pointed out, faultrupture locations can be caiculated, fundamental parameters äs dissipated energy and seismic moment are estimated and seem realistic. The three-dimensional aspect in modeling is pointed out äs very critical. However, at the present time, no calibration can be demonstrated. As well, impact of failure along the structures on mining areas are impossible to estimate on a modeling point ofview. Therefore, this methodology of three-dimensional modeling by the distinct element method appear äs a method to be calibratcd äs closeiy äs possibie (figure l O): -above all with seismic results from the setüed network in 1991, giving presently locations and daüy seismic energy distributions. Back analysis of selected tremors (location, energy, radiation pattern,...) compared with modeling could provide very interesting results. The seismic network is still undergoing improvements to locate more accurately, particularly in depth, and to gain understanding of the focal mechanisms involved in rockburst occurences. Association of seismic analysis and dynamic modeling is expected to come out äs very promising. In every case, this research directiy benefits the mine by providing a method able to quantify numerous potential fault-slip Problems, 'or qualify rockburst mechanism.

  fig. 2-kinds of rockbursts

  fig.3-schematic description ofthe damages flg. 4 -3DEC scheme caiculations (afterHart, [1988])

  fig.8: parameter M at each step for structures A & I

  fig. 10 interacting fields to exchange data -with insitu observations, i.e. detailed, geological survey of the faults and their characteristics, of the local disturbances ofthecoalseam, and othersparticular Underground conditions.

table l -

 l elasto-plastic parameters of the model

	table l	structural features	rock matrix MPa
	stiffnesses	kn,ks=10000MPa/m	K,G=13333,8000
	M.C. Parameters	Fric=35' > ,Coh=0, Rt=0
		MPa
		table 2 -input Stresses
	table 2
	value MPa

ACKNOWLEDGEMENT

we greatly appreciate the help ofDr Fairhurst and Dr Piguet, who made that work possibie.


J U Bigourd

INERIS J M Dassonville

THE GROOVE MORTAR, A REUABLE METHOD FOR TESTING THE FLAMEPROOFNESS OF CERTAIN EXPLOSIVES

For more than sixty years, cylindrical steel cannons (long mortar and short mortar) have been used in France to assess the safety of explosives with regard to firedamp. These testing methods are considered expensive mainly because a large number of shots are needed äs a result of the low level of reproducibility. Also, using them with certain explosives, such äs slurries and emulsions, is somewhat problematical. Taking into aecount the Polish expertise, CERCHAR (now INERIS) developed m the sixties a testing method in which the Charge, placed in a steel groove mortar, is isolated from the methane-air mixture by an asbestos-cement plate. This method seems more attractive than those traditionally used, especially when it comes to distinguishing rock explosives from the more powerfui pennitted explosives.

INTRODUCTION

The first part of this document examines the factors which encouraged us to look for a new method for testing the flameproofness of explosives. The following parts indicate how the method was developed, together with a description of the present tests and a summary of the results obtained so far.

Explosives in French coal mines

Under French regulations, explosives designed for use in Underground coal mines can be divided into three groups : -group l (in French "explosife rocher") : rock explosives, which are only allowed in the least dangerous driftings, -group 2 ("explosifs couche"): coal getting explosives, which are mainly used for blasting operations with instantaneous firing, in main Ventilation areas, -group 3 ("explosifs couche ameliores"): improved coal getting explosives which can be used with short delay detonatois, in the most dangerous working places where blasting is pennitted.

All these explosives must have a good detonation ability and a low production of toxic gases when they detonate. For group 3 explosives, a low deflagrability is also required. With regard to safety in the presence of firedamp or coal dust, there is no requirement for group l explosives, unlike group 2 and 3 explosives for which tests must be carried out. For group 2, the approval conditions are not very severe, the main aim being to avoid ignition of a coal dust deposit äs the result of a blown out shot. The approval conditions are more stringent for group 3. First and foremost, it means preventing the ignition of a cloud of coal dust (in an atmosphere which can contain a small amount of methane) in the case of exposure during delayed blasting. Consequently, it was believed that by carrying out a reasonable number of shots, the characteristics of a given explosive could be determined by its Charge limit (maximum Charge which does not cause ignition in a fixed number of tests, usually five). When compared with what really happens, there is obviously no basis for the notion of Charge limit. What is more, the state of wear of the mortar has a certain influence on the results, leading to a lack of reproducibility since it is not financially possible to change the mortar for each test. Thus, the ignitions observed with the smallest charges were mostly obtained when the mortar was new (first ten shots).

-Figure l -Long mortar; probability (P) of firedamp ignition äs a fünction of the length (L) of the Charge; (I) old theory; (R) reality.

Tests versus reality

Originally, according to [START_REF] Coward | Research on the mcendivity of Coal-Mine Explosives : A Review[END_REF], the steel mortar was chosen, in preference to blocks ofrock or coal, for simplidty in preparation and with the reasonable hope of greater reproducibility of results. The tests carried out in the different experimental mines [START_REF] Nagy | Hazards of Cut-off Explosive Charges in Multiple Blasting of Coal[END_REF][START_REF] Meerbach | Versuche zur Zündung von Schlagwettern unter Grubenbedingungen[END_REF][START_REF] Cybulski | Investigations on safety of explosives in presence of methane and coal dust, in coal and sandstone äs well äs in steel mortars[END_REF] enables the official tests to be compared with the practical conditions. From these tests, it emerges that, generally speaking, the use of steel makes the tests more severe but, in certain cases, would seem to introduce artifacts. For example, it has been observed that Variation in a parameter can have the opposite effect in a test and under the corresponding practical conditions.

However, the Polish slot mortar test, where a layer of clay separates the explosive Charge from the methane and air mixture is an exception. It gives results in the vicinity of those obtained under sünilar conditions, for rock and coal, in the Barbara experimental mine [START_REF] Cybulski | Investigations on safety of explosives in presence of methane and coal dust, in coal and sandstone äs well äs in steel mortars[END_REF]. In 1965, we started developing a similar test which would enable us to do away with long and short mortar for the approval of group2 explosives. This is the test which is presented here.

DEVELOPMENT OF THE GROOVE MORTAR TEST

The Polish slot mortar is a cylindrical steel mortar, a sector of which is removed over a certain length. It is therefore relatively fragile when powerfül explosives are fired. To obtain a more resistant apparatus (figure 2), we milled a shallow, trough-shaped groove on one side of a block of steel. This groove, designed for 30mm diameter cartridges, has a depth of 40mm and a width of 50mm (figure 2). Provided the Charge is correetly placed in the groove section using a suitable device, a free space of at least 10mm can be left between the Charge and the steel, in order to reduce the wear of the mortar.

By covering the groove with a plate of material comparable to rock, we can simulate the case of a mine hole whose lateral confinement is reduced during firing of the round. To encourage lateral expansion of the gas rather than axial expansion (figure 3), the groove is filled with clay tamping (humidity 15 %), over a length of 100mm, on either side of the Charge (m a long mortar shot, this amount of clay, used äs stemming, largely reduces the risk of firedamp ignition).

We conducted several tests m which the groove was covered with a clay plate. We obtained comparable results to those of the Barbara experimental mine. With a group 2 explosive, a thickness of 10 to 20mm of clay was sufficient to prevent ignition of the firedamp when the length of the Charge was between 200 and 500mm. However, making the clay plates was difficult and expensive. We decided to replace the clay by asbestos-cement, a material which is commercially available in plates of various thicknesses and which simulates rock fairly well. Both materials have sünilar densities but behave differently when an explosive Charge detonates into contact with them. X-ray flashes show that blasting of asbestos-cement is faster than that of clay, for the same thickness of material.

Since the shot fümes are not contained äs long, it can be expected that the asbestos-cement will be a little less effective than the clay in preventing firedamp ignition.

In our first tests, we used 80mm wide asbestos-cement plates.' Firedamp ignition occurred for the group2 explosives, regardless of the thickness of the asbestos-cement plate (6 to 60mm), even for small charges (100mm long). l/25th second photographs, taken with an argon flash, showed that the blasting fumes get between the asbestos-cement and the steel, causing a large amount of luminosity on either side of the plate. This probably explains the ignitions observed. By increasing the width of the plates, we then obtained results which had a closer resemblance to those of the Barbara experimental mine. The width of the asbestos-cement plate is now fixed at 200mm. We also make sure that the plate is in dose contact with the steel. When the edges of the groove are too uneven for this to be possible, we grind the surfäce of the steel block or we change the groove.

We varied the thickness of the asbestoscement plate so äs to detemüne the minimum thickness which does not result in ignition after lOshots, for a fixed Charge length. These tests were carried out with different batches of a group 2 explosives. For these batches, the ballistic mortar work ranged from 78 to 87% of the picric acid strength. For charges with a length of 300mm or more, the minimum thickness varied considerably from one batch to another, increasing with the ballistic mortar power measured. For charges of 100 or 200mm, the results were more regulär -a minimum thickness of 6 or 10mm. We chose to impose an asbestos-cement plate thickness of 6mm and vary the length of the Charge from now on m order to determine the Charge limit.

-Figures 2 and3

THE CURRENT GROOVE MORTAR TEST

Our firedamp test gallery consists of a 12m long steel outer wall with a diameter of 2m, closed at one end and open at the other. A sheet of paper, inserted in a slot located 5m from the closed end divides the gallery into two sections -the Bring chamber between the paper and the closed end, and the expansion chamber. The firing chamber, where the an-and methane mixtures are produced, has a volume of lim 3 . The sides are protected with steel armour at the closed end and with an inside concrete coating on the cylindrical part.

At present, the groove mortar consists of a steel block attached to a cart. The block is 1.2m long and has a square section with a side of 0.35m. It has a groove on each of the four main faces, so that its Position can be changed by rotating the block one quarter turn. The groove used is that facing the closed end of the gallery when the mortar is positioned in the firing chamber (figure 4). During the test, the distance between the groove and the closed end ofthe gallery is 2,5m.

The asbestos-cement plate'used is 6mm thick, 200mm wide and at least 200mm longer than the Charge being tested. Every 100mm along its main axis, the plate has two small holes (2mm diameter), 30mm apart and symmetrical around the axis. A thin wire device (1mm diameter) through the two holes is used to attach the Charge (cartridge diameter 30mm) to the plate. The wire device has two 10mm long lugs to ensurc that the Charge will be centred in the groove.

The groove is filled with clay (15% humidity) along 100mm on either side of the füture location of the Charge. The charge-plate assembly is then placed and held placed with adhesive tape. After the methane-air mixture has been produced, firing takes place.

The length of the Charge can take a series of values in an approximately geometric Progression ( ..., 80, 100, 125, 160, 200mm, ...). It is varied from one test to another, in order to determine the Charge limit. The Charge limit has a length equal to or greater than 125mm for group2 explosives, and less than or equal to 50mm for typical group l explosives. The Charge limit (100mm) is between the two previous groups in the case of the emulsion h which is not especially designed to be a safety explosive. These results seem fairly well related to the power of the explosives determined by ballistic mortar.

CONCLUSION

Determining the Charge limit using groove mortar with a 6mm thick asbestos-cement plate is a method which could replace our traditional long and short mortar methods for evaluating the flameproofness of group 2 explosives in the presence of firedamp. It has the following advantages over traditional methods:

-it can be considered more realistic in the light of information obtained äs the result of tests carried out in various experimental mines; -it is sufficiently reproducible and enables group2 explosives to be distinguished from classical group l explosives with only a small number of tests; -its use with powerfül explosives of the emulsion type pose little problem, since it uses a relatively resistant apparatus; -äs a result of the last two points, it is less expensive.

  -Groove mortar; cross and longitudinal sections ; (a) steel block; (b) clay tamping; (c) Charge; (d) detonator; (e) lug; (f) iron wire device; (p) asbestos-cement plate; (r) groove.

  Test gallery with groove mortar in test position; (a) steel block; (b) steel plating; (c) cart; (p) sheet ofpaper; (r) concrete coating; (v) steel shell.

Table l

 l

	gives the Charge limit obtained
	for various group l and 2 explosives, äs well äs the density of the explosive and
	its ballistic mortar strength (in % of the
	picric acid work). Explosives a, b, c, d and A are TNT sensitised explosives.
	Explosives e, f, g and B are NG
	explosives. Explosive h is an emulsion. In group 2, explosives A and B contain approximately 20 % sodium Chloride.


G Rose

IMPROVING FIREDAMP METERS PERFORMANCES BY MEANS OF ELECTRONIC TECHNIQUES

The sensor used in French mines to detect firedamp (methane) in airways is a platinum filament catalytic sensor. Its perfonnance characteristics have been constantly improved by new conditioning techniques: heating filaments functionning at high temperature, signal processing. In this document, we describe the operadon of the sensor and the conditioning techniques successively used in different mine gas detectors, stressing for each the improved performance obtained. In conclusion, we will show that the last technique recendy developed broadens the field of use of the sensor to industrial and domestic explosimetry.

INTRODUCTION

Sensors used in explosimeters for the detection of explosive gases are generally catalytic. The most widespread are catalytic pellets.

A metallic oxidation catalyst is heated to a sufficiendy high temperature to oxidize the gas. The oxidation reacüon increases the temperature of the catalyst which increases its electrical resistance, whose amplitude is proportional to the gas concentration.

In the 1950s we developed a sensor of this type for the mining industry. It was based on a platinum filament and destined to detect methane, responsible for mine explosions.

The filament is distinguished from the catalytic pellets by its operating temperature: 1000°C instead of 600°C, which confers on it the advantage of being relatively insensitive to fouling and catalyst poisoning. The disadvantage, however, is rapid wear, preventing it from operating permanentiy.

The measurement technique thus involves energizing the sensor just long enough to make the measurement and then tuming it off for äs long äs the application pennits. The first instruments of this time could run for 100,000 measurements, or about 9 months at the rhythm of one measurement every 4 minutes.

By improving the filament heating technique, the measurement capacity has been increased considerably: 10,000,000 measurements with the last methanemeter rccenüy markeied, or a lifetime of 1.5 years at the practically continuous measurement rhythm of once every 5 seconds.

Endowed with these performance characteristics, the platinum filament sensor is today a reliable solution for explosimetry, in particular in environments polluted with certain catalyst poisons such äs sulfur-containing products, which other catalytic sensors resist poorly.

In addition, the technique involving brief heating pulses, applicable only to filament sensors since it requires a very low thermal inertia of the sensor, has enabled new procedures to be developed for the identification of certain combustible gases or vapors and for the measurement of methane concentrations up to 100% v/v. It is composed of two Spiral filaments made of pure platinum: a detector filament and a compensating filament, both 80 m in diameter but of unequal lengths.

DESCRIPTION OF THE

The filaments are enclosed inside a hemispherical chamber made of sintered metal, enabling the natural diffusion of gases. They are welded to massive high thennal inertia antennas, that emerge from the chamber through isolated ports.

An activated charcoal filter inside the chamber surrounds the filaments and protects them from catalyst poisons, in particular silicones.

OPERATION

Petector filament

During the measurement period, a current I passes through the filament, raising its temperature in the center to 1000°C by the Joule effect.

Its electrical resistance Rd is a function of temperature T, which depends on I and on the quandäes ofoxidized gases.

The temperature of the filament is not the same along its entire length because of the cooling caused by the metallic Supports.

Catalyüc activity occurs on the central spires, whose temperature is higher than 800°C. The detector filament is appropriately treated to eliminate impurities in platinum.

Its temperature is determined in order for catalytic activity to be maximal at low methane concentrations, lower than the L.I.E., and to bum off products that could foul its surface, thereby reducing the acrive sites on which the oxidarion occurs.

-2-

Compensatin? filament

This filament does not undergo a purifying treatment and is heated at a lower temperature, about 600°C on the central spires. At this temperature, catalydc activity Starts at the stoichiometric concentration of 10% v/v for methane.

The compensator has a dual role: it compensates the thermal effects that interfere with detector measurement (temperature and thermal conductivity of the ambient medium), and detects high levels to remove any doubts concerning detector measurement.

Rc (600°C) = 0.73 . Rc(0°C) = 0.29 signal classical sensor: without "removal of doubt" %v/v %v/v bl l l l l l l70

10 100 S(2%) = S(75%) signal CHARACTERISTICS Lifetime
Sensor lifetime is that of the detector filament. Platinum wear begins at 600°C and is particulariy high around the operating temperature of the filament (at temperatures dose to 1000°C, the rate of wear of the platinum filament is proportional to the cube of temperature).

It is thus desirable to control this temperature throughout filament life, in particular when there is a supply of additional heat resulting from the oxidation of a gas.

Filament wear, greater in the central zone, results in a reduction in the cross-section of the wire, increasing resistance when cold Rd(0°C) and when hot Rd(T). This increases the explosimeter readings by a constant quantity and thus requires frequent recalibration of apparatus zero (reading m pure air).

Sensor lifetime thus depends on the duration of the measurement and on filament temperature during the measurement.

The various heating techniques used in several generadons of methane detectors in order to increase lifetime have attempted to reduce the duration of the measurement and to control temperature.

Response dme of the methane detector

HEÄTING TECHNIQUES AND PERFORMANCES

A number of different techniques have been used since the creation of the sensor, äs developments in electronic components made them possible. We will describe them in the chronological order of development of the methane detectors that used them.

Constant intensitv heäting : remote mine ga.s detector CTT63/4Q fi ™ mine gas oetector

• Signal = 50mV x %v/v CH4

• Theat = 3 s

• Wmeas = 2.9 J

• life time:

. in air = 100,000 measurements . in 2% CH4 = 30,000 measurements The potential differences at the terminals of the compensator and detector filaments are held constanL The potential difference at the tennmals of the two filaments mounted in series is adjusted to 1.300 V.

A potential limiting circuit prevents Ua from exceeding 0.760 V by shunting a current (Imeas) when Rd increases

The decreased current in the detector attenuates the temperature increase caused by gas oxidation or filament wear. The effect is more pronounced than in the case of the constant bridge voltage set-up.

Increased measurement capacity is partially due to the reduced heating time, brought down to 1.3 seconds (in order to enable the measurement to be read, a circuit keeps the measurement in memory during filament extinction).

The intensity of the heating current, high at the beginning since the filaments are cold, enables operaüng temperature to be reached in about l second.

The measurement signal is memorized during the last 300 ms. The increase to operating temperature occurs very rapidly by a constant high intensity current When this temperature is reached, a circuit maintains the resistance at Rd (1000°C) regardless of gas concentration, by acting on current intensity : if addidonal heat are supplied by the oxidation of a gas, the circuit decreases the Joule effect heat by reducing the current.

Rd(T) = Rl x R3/R2

Temperature rise of the filament lasts 0. When Rd(T) = 1.03 ß, the Controller reduces intensity in order to maintain Rd at this value.

Variations of the plateau potentiel Ud is the measurement signal. The measurement is memorized before the current is cut This heating technique confers a very long lifetime on the sensor : 10,000,000 measurements and independent of the quantities of gas oxidized by the filament. This is due to 3 factors :

-reduced heating time, -temperature control during measurements, -slight decrease in filament temperature äs it wears: with increasing wear, Rd(0°C) increases. Since Rd(T) remains constant (1.03 W) filament temperature decreases.

Wear is independent of the gas history of the filament. In this case, it depends only on the number of measurements, which enables zero drift to be simulated in order to carryout an automatic correction and thus considerably reduce maintenance operaüons.

Consumption is very low äs a result of the short heating time and the suppression of the compensating filament Comment: the compensating filament can be replaced by a low consumption temperature sensor (thermistor, semiconductors, etc.). Although changes in relative humidity are not compensated, their influence on measurements remains negligible, since they are lower than 0.1% v/v CH4 at 80% R.H. cadmium battery is sufficient. -Automatic zero correction: every month.

The "removal of doubt" at high methane concentrations, which was assured by the compensator and made indispensable by the low measurement rhythms in the initial methane detectors, ean be replaced by a "locking" alarm circuit, äs in condnuous measurement explosimeters.

A practically condnuous measurement rythm of 5 seconds leads to a consumpdon of 0.024 Joule, i.e. a mean current of 0.07 A.

The device requires a power supply higher than 1.2 V. A 2-element 2.5 V nickel-

Summary table

The

Conclusion and perspectives

The properties of the platinum filament sensor, designed 30 years ago for mining applications, are today optimally used äs a result of the perfected conditioning of its power supply. At a similar response time, its electrical consumption is lower than that of low consumpdon catalytic pellets.

Its high operating temperature effectively protects it against certain catalyst poisons.

Its field of application now extends to industrial and domesäc gas detection. The technique of heating by short pulses, specific to the filament with low thermal inertia, enables the pnnciple faults of catalytic oxidaäon to be overcome, with several improvements : measurement ränge Umited to the stoichiometric value ofthe airgas mixture, and absence of selecdvity.

Examples:

-0-100% v/v mefhane detector The heating pulse is decomposed into two temperature plateaus:

-Ist plateau at T < 400°C: the filament operates äs a hot wire detector and fumishes measurements of concentrations higher than stoichiometry,

-2nd plateau at 1000°C: the filament gives the measurement of concentrations lower than stoichiometry.

The zero of the hot wire detector is done automatically each üme the measurement is lower than the L.I.E.

  SENSOR 25eme Conference Internationale des Instituts de Recherches sur la Securite dans les Mines, Pretoria, 13-17 septembre 1993, vol. l, p. 31-39.

  )=Rd(0°)x (l+kt°) t°=f(I;%v/vCH4) 1°= mean temperature of the filament tmean = l / L } t (dl) 0 Rd = resistance of detector filament at t°k =0.004 for the platinum Rd(0°) = resistance at 0° = p x L / S p = resisüvity; L = length ; S = section Rd(0°) = 0.35 ß Rd(1000°) = 1.03 ß -Rd(1000°) = 1000° on the central spire.

  response time = Tsens + Text cas 2-response time = Tsens The response time of a methane detector or an explosimeter is the time it takes to indicate 90% of the value of a concentration of gas G. The response time of the Instrument depends on: -the response time of the sensor (Tsens), which is the time the gas takes to fill in the combustion chamber, -the extinction time (Text) separaüng two consecudve measurements, -the occurrence of the measurement in relaäon to the gas concentradon. If Text > Tsens then: Tsens < response tinae < Tsens + Text Electric consumpfion Wmeas = energy consumed during the measurement = (RdxId^RcxIc^x Theat Pmeas = power dissipated during the measurement = Rd x l2 Pmean = mean power = Pmeas x Theat/(Theat+Text) Mean power takes into account the extinction time of the filament. It determines battery autonomy for portable apparatus. Theat = filament heäting üme = duration of measurement Text = durarion of filament extinction

  temnerature ri.se and constant detector filament resistance : air and gaz) = cte = 1.03Q the compensator filament is not used

  Characteristics of the GTM90C mine gas detector using this technique • -Measurement rhythm: 5 s, automancaüy increases to 2.5 s if the concentradon increases, -Sensor lifedme: 19 months, -Response dme: 15 to 18 s, -Response dme for the alarm to an explosive concentradon : < 5 s, -Consumpdon : 0.09 A (0.02 A for the electronic circuits),

  table summarizes progress in the characterisdcs and perfonnance of the different techniques.

			d.d.p. at terminals detector filament (Volt)	intensity through the filament (Ampere)	filament resistance (Ohm)	consumpdon during the measurement (Joule)	lifeume or capacity of measurements
	constant intensity	air 2%	0.76 0.86	0.74 0.74	1.03 1.15	2.9	100,000 30,000
	air bndge ---constant	0.76	0.74	1.03	2.9	250,000
	voltage CH4 2%	0.80	0.72	1.12		100,000
	constant alr pardal --voltage CH4 2%	0.76 0.76	0.74 0.71	1.03 1.08	1.25	2,000,000 1,000.000
	air resistance CH4 constant 2%	0.76 0.72	0.74 0.70	1.03 1.03	0.36	10,000,000 10,000,000
				-8-

  The catalytic acüvity of a combustible gas or vapor äs a function of oxidation temperature follows a law specific to each species. The combustible product can be identified by measurements m ade at different temperature plateaus. Zeros at the 3 temperatures are done automatically when the air is pure.

		-Idendficaüon ofthe gas
		®	tO 0.2 0.3 0.4
			^------t
			heat
		INERIS (Institut National de
		1'Environnement Industriel et des Risques)
		develops these new techniques involving
		signal processing and algorithm execution,
		and is continuing work on the platinum
		filament sensor in order to improve its
		perform ance characteristics ; thus, a new
		smaller one, to reduce response time, is
		being developed.
		The sensor and the methane detectors
		described in this document are
		manufactured by Oldham-France wider an
		INERIS license.
	t0	0.2 0.4 s
		theat