Method and apparatus for fracturation detection

A method and apparatus for detecting cracks or other separations deep inside massive rock-like structures and assessing the integrity and safety of mine roofs and the like, wherein at least one borehole is drilled into the rock and a negative fluid pressure is applied thereto. The permeability of the rock at selected intervals along the borehole is measured and thereby the presence or absence of fractures in the rock is detected. The extension of the fracture to an adjoining hole may be determined by a pressure respond test in the adjoining hole.

FIELD OF INVENTION 
This invention relates to a method and apparatus for detecting cracks in 
massive solid structures. More particularly this invention relates to 
methods and apparatus for detecting cracks, fractures or separated rock 
blocks in rock masses such as mine roofs and walls or in concrete 
structures or the like. 
BACKGROUND OF THE INVENTION 
Mine accident investigation statistics reveal that a surprisingly large 
number of fatal accidents occur every year through unpredictable roof 
falls, even in mines where all presently applicable safety regulations 
stipulated by the law are rigidly and meticulously enforced. Such 
accidents are routinely classified as "Act of God" because heretofore 
there has been no way for mining safety engineers to detect and monitor 
the development of dangerous, unstable rock conditions underground. 
Heretofore, the only methods to check the roof conditions have been (a) 
observation of the tensile strength increment on the rock bolts which 
indicates stress build-up in the roof and (b) "sounding" the rock by 
tapping it with a scaling bar. It will be appreciated, however, that 
neither of these methods is infallible because (a) tensile stress build-up 
may not occur because the rock bolts themselves may move with the sagging 
rock block and (b) the developed separations are initially very small and 
can be located very deeply within the rock mass so that tapping will fail 
to reveal any unsoundness. 
BRIEF DESCRIPTION OF THE INVENTION 
It is an object of the present invention to provide a simple and 
inexpensive method for detecting the presence of potentially dangerous 
cracks in massive structures such as the roof of an underground mine. 
It is another object of the present invention to provide apparatus for use 
in the aforesaid method. 
Thus, by one aspect of the invention there is provided a method for 
detecting fractures and determining the degree of fracturation in a rock 
mass or concrete of known matrix permeability having at least one borehole 
therein, comprising (a) pressure sealing a selected length of a selected 
borehole; (b) applying a negative fluid pressure to said borehole, and (c) 
measuring the fluid transmitting capacity of said rock through selected 
intervals along said borehole relative to matrix permeability. 
By another aspect of this invention there is provided apparatus for 
detecting fractures and determining the degree of fracturation in a rock 
mass or concrete of known matrix permeability having at least one borehole 
therein, comprising: 
(a) pressure sealing means to seal a selected interval along a borehole in 
a self-supporting rock mass; 
(b) means to apply a negative fluid pressure to said sealed selected 
interval; 
(c) means to measure fluid transporting capacity through said sealed 
selected interval relative to the matrix permeability of said rock mass.

DETAILED DESCRIPTION OF INVENTION 
Rocks encountered in underground mining operations are generally considered 
to be heterogeneous because they contain, in varying degrees, geological 
fractures, fissures, joints, solution channels, cracks, faults and other 
mechanical discontinuities which create interfaces in the rock mass. Bond 
of tensile strength across these interfaces or partings can vary from zero 
to a value equal to or greater than the tensile strength of the host rock. 
During excavation of a stope, time-dependent deformations take place which 
tend to break fractures, joints and fault bonds creating instability in 
the roof. If the bond or tensile strength across the interface is very low 
the roof rock over the stope can become detached from the rock above to 
form a slab loaded only by its own weight (FIG. 1a) or additionally by 
thinner less rigid slabs above (FIG. 1b). This slab is known as the 
immediate roof and can be considered as a separate unit, the moments, 
shears and stresses of which can be calculated by beam theory. 
The equations for a horizontal roof single slab are: 
##EQU1## 
and the formula to calculate the thickness of a self-supporting beam is 
##EQU2## 
where T.sub.max =maximum shear stress 
.theta.max=maximum tensile stress 
.gamma.=unit weight of rock 
S=span of stope 
t=thickness of immediate roof 
The relationship between the thickness of a self-supporting beam and the 
span of a stope is well known and is illustrated in graphical form in FIG. 
2. It can be seen that, for a horizontal roof slab, a relatively thin 
immediate roof slab can be self-supporting and provided rock bolts are 
installed shortly after excavation, to secure the immediate roof to the 
main rock mass, the development of an immediately dangerous roof condition 
is very unlikely. 
However, in the presence of joints and slip planes a completely separate 
rock block can be developed which is very weakly bonded to the main rock 
mass, and which is sometimes known as a "coffin roof" because of the 
dangers inherent therein. If the support under such a rock block is 
gradually removed by the mining operations, serious roof failure becomes 
imminent, depending on the orientation of the interfaces and the direction 
of stoping. Three different conditions can be developed in either a 
parallel or perpendicular plane relative to the direction of excavation: 
(a) loose rock block supported at both ends; 
(b) loose rock block supported at one end only; and 
(c) loose rock block without support. 
The loose rock block in FIG. 3a is supported at both ends, acts as a self 
supporting beam clamped at both ends, and is inherently safe. If however 
the span of the stope is widened so that the planes of the interface tend 
toward intersection with the side walls of the stope, the support for the 
coffin roof is removed and it becomes unstable, as illustrated in FIG. 3b. 
Similarly the loose rock block shown in FIG. 4a acts as a gravity loaded 
cantilever beam and is inherently safe. FIG. 4b, however, illustrates what 
happens as the support is removed as the stope advances. The weight of the 
unsupported block gradually breaks the bond at the interface and roof 
collapse occurs. 
Deterioration of a mine roof is generally a gradual process and actual 
collapse usually occurs several days or even weeks after initiation of the 
separation. 
It is also known that the reduction of bond or tensile strength is always 
accompanied by the development of fine separations or cracks. Such cracks 
significantly increase the gas transmitting capacity of the rock mass. 
Media that allow the passage of a fluid through them are said to be 
permeable. If the fluid movement takes place only in the naturally 
occurring pores and fractures and separation is absent, the permeability 
is referred to as "matrix permeability" and it is important to recognize 
that this represents permeability before fracturation. Absolute 
permeability expresses the fluid transmitting capacity of the medium when 
the pore space is 100 percent saturated with the fluid flowing in the 
medium. Absolute permeability of one darcy may be defined as that 
permeability which allows the flow of 1 cc per second of fluid of 1 
centipoise viscosity through an area of 1 sq cm under a pressure gradient 
of 1 atm. per centimeter when the flow condition is viscous. It is known 
that the matrix permeability of unfractured rock normally encountered in 
underground mining operations is about 0.0001-0.001 darcies. As indicated 
in FIG. 6, however, rock with a 0.1 mm fracture through it, has a 
permeability of about 837 darcies, i.e. 800,000-8,000,000 times greater 
than the unfractured rock. 
Rock/roof bolting is an accepted technique for reinforcing roofs in all 
types of rock. The function of the rock bolt is to bind together the 
stratified roof and suspend it by a bolt anchored in the self-supporting 
rock above it. Holes are drilled in the roof, at appropriate intervals, 
into the self-supporting rock mass and long steel bolts are inserted 
therein and anchored in the solid rock mass. A plate can then be secured 
to the free, lower end of the bolt and tensioned as desired to support the 
roof. It is convenient, therefore, that conventional rock bolt holes, 3-5 
meters in length, be employed in the present invention according to which 
one hole is selected as the vacuum hole 1 while one or more other holes 
may be designated observation holes 2 (FIG. 5). A straddle packer 4 
(described in more detail hereinafter) is inserted into vacuum hole 1 and 
a negative (vacuum) pressure is applied thereto by a vacuum pump 9 
connected via line 10. It is of course well known that a fractured zone 
develops immediately adjacent the stope due to blasting and it is, 
therefore, important that the packer 4 should be installed beyond this 
zone so that the isolated zone is in a theoretically crack-free zone. The 
packer may be of the single or straddle variety. The single packer may be 
used to investigate the total length of a hole, but in order to locate and 
isolate a fracture at a particular position along the length of the hole, 
it is necessary to use a straddle packer which is also suitable for use in 
measuring the static, i.e. the "rock face" pressure directly FIGS. 8, 9, 
10 and 11 illustrate a typical straddel packer suitable for the purposes 
of the present invention. A novel feature of the packer is that it uses 
dual flow lines to measure the dynamic and static i.e. the rock face 
pressure simultaneously with the flow rate measurement. The accurate rock 
face pressure measurement is the basic requirement for fracture 
evaluation. The packer assembly 24 is run in the test hole 25 by using the 
proper number of extension rods 26 and extension tubings 27. Setting tool 
7 (in FIG. 5) or 85 (in FIG. 8) is used to set the packer at the selected 
location. The packer 24 is kept at the selected position by using the 
holder bolt 28 and plate 29 and stand 6 (in FIG. 5). By turning the 
setting wheel 30 and moving the setting pipe 31 upwards, rubber elements 
32 are compressed by the forward movement of the extension tubings 27, 
cross-over tubing connector 33, protector pipe 34 and spacer 35. Peg 36 
prevents the turning of the extension tubing 27 during installation. After 
the deformation of the packers 37 (81, 82 in FIGS. 8 and 9) the 
investigation interval 38 (or 86 in FIG. 4) is completely isolated. Air 
flow is created by vacuum pump 9 (in FIG. 5). Air from investigated 
interval 38 enters into the hollow section of the base rod 39 through 
spacer 40 and base rod port holes 41 (87 in FIG. 9). Air is removed from 
the hollow section of the base rod 39 through flow line connection 42 and 
tubing flow line 43. The separate measuring line 44 is connected to the 
hollow section of the base rod 39 by the static line connection 45. The 
static line connection 45 is placed perpendicularly to the hollow section 
of the base rod 39 where no air movement takes place in order to measure 
the "rock face" pressure accurately. In order to release the packer it is 
necessary to turn the setting wheel 30 downward. It is mandatory to check 
the roof condition of the working face of a stope as close as possible 
where blasting is performed regularly. For this reason a packer which can 
not be damaged by blasting was developed. The packer uses dual flow lines 
to measure the dynamic and static i.e. the rock face pressure 
simultaneously with the flow rate (FIGS. 8, 9 and 11). The packer is 
placed at the selected section in the hole by using outside 46 and inside 
47 extension pipes. The packer is set by compressing the rubber elements 
48 by turning clockwise the coupling nut 49 with the turning head 50 
fastened to the first outside extension pipe 46. During this operation the 
packer is held in position by the pipe holder 51 fastened to the first 
inside extension pipe 47 which prevents the turning of the packer. The 
deformation of the rubber elements 48 isolate the investigation interval 
52 between the elements. After the installation of the packer, the outside 
46 and inside 47 extension pipes with the turning head 50 and the pipe 
holder 51 are removed. The only exposed parts of the packer assembly are 
the tygon.RTM. flow line 58 and the tygon.RTM. static pressure measureing 
line 60. The air flow is induced by a vacuum pump. Air enters in the flow 
pipe 53 via spacer port holes 54 drilled in the spacer pipe 55 and base 
pipe port holes 56 drilled in the base pipe 57. The flow pipe 53 is 
connected to the vacuum pump by the tygon.RTM. flow line 58. The static 
i.e. the rock face pressure, is measured in the annulus 59 where no air 
flow takes place. The annulus 59 is connected to the static manometer 18 
(FIG. 1) via tygon.RTM.static pressure measuring line 60 through port hole 
61. The packer is released and removed from the hole by using the outside 
46 and inside 47 extension pipes with turning head 50 and pipe holder 51. 
In this case the coupling nut 49 must be turned counterclockwise. As shown 
in FIG. 5, vacuum hole 1 and observation holes 2 drilled in the roof, 
intercepted a fracture 3. Straddle packer unit 4 is installed at various 
depths in the vacuum hole by using the extension rod assembly 5 and stand 
6. The straddle packer 4 is set by turning the setting assembly 7 and 
isolate a section of the hole between the rubber elements 8. Then vacuum 
is created by the vacuum pump 9 through the flow line 10. The magnitude of 
the vacuum is measured by mercury manometer 11, water manometer 12 or 
vacuum gauge 13 by using static line 14. By moving the straddle packer in 
the hole and repeating the test the fracture 3 present in the rock mass is 
intercepted. In case of very small fractures (microfracture) the flow rate 
is very low and cannot be measured by commercially available flowmeters. 
In this case the flow rate is established by observing the absolute 
pressure change during unit time in a closed system which consists of the 
air volume between the rubber sealing elements of the packer, connection 
tubings, fittings manometer etc. The flow rate is calculated by using the 
following formula: 
##EQU3## 
where: Q--flow rate cc/min at 1 atm and 60.degree. F. temp. 
H.sub.1, H.sub.2 --vacuum in Torr at the start and end of the observation 
period respectively 
t--length of observation period--(minutes) 
T.sub.f --flowing air temperature (F..degree.) V--volume of closed system 
(cc). 
To calculate the reservoir engineering parameters of the fracture the 
following formulas apply. 
##EQU4## 
where: K=permeability (md) 
T=transmissibility (darcy--feet) 
H.sub.x =pressure loss (Torr/min) 
P.sub.e, P.sub.w =pressure at radius r.sub.e and r.sub.w (psia) 
W.sub.mm =fracture size (mm) 
h=investigation interval (25 cm) 
d=packer diameter (1.5 in.) 
v=vol. of closed system (500 cc). 
In the case of larger fractures the flow rate may be measured by using area 
type flowmeter (rotameter) 15 or horizontal flowmeter 16 and flow line 10. 
By the application of the exhaust assembly 17 described in more detail 
hereinafter, the exhaust air is measured with the flowmeters at 
atmospheric condition instead of the flow rate under negative pressure. 
The prevailing negative pressure between the rubber elements (generally 
called "rock face pressure") which creates the air flow is measured by the 
mercury manometer 18 or water manometer 12 or vacuum gauge 13 by using the 
static line 14 in which no flow takes place. To calculate the reservoir 
engineering parameters of the fractures under these conditions, the 
following equations apply: 
##EQU5## 
where: Kf=permeability (darcies) 
T=transmissibility--(darcy-feet) 
Q=flow rate (1/min) 
P.sub.e, P.sub.w =pressure at radius r.sub.e and r.sub.w (psia) 
W.sub.mm =fracture size (mm) 
W.sub.in =fracture size (in.) 
d=packer diameter (1.5 in.). 
In order to establish the geological parameters i.e. strike, dip and areal 
extension of the fracture detected in hole 1, additional "observation 
holes" 2 may also be tested in conjunction with vacuum hole 1. A straddle 
packer 19 is inserted into each hole 2 in the same manner as in hole 1 and 
a test is conducted at selected intervals therealong until the fracture 3 
intercepting hole 2 is isolated by straddle packer 19. The installation 
and setting the straddle packer in the observation hole 2 are identical 
with the procedure used in the vacuum hole 1. The continuity of a fracture 
between holes is established by observing the effect of the created vacuum 
in the vacuum hole 1 with the straddle packer 19 installed in the 
observation hole 2. The response is observed by mercury manometer 20 or 
water manometer 21, or inclined manometer 22, or vacuum gauge 23. A 
positive response indicates direct pneumatic connection between holes 
through the fracture. Pneumatic connection indicates that the tensile 
strength at the interface is zero. The utilization of a selected number of 
observation holes will determine the geometry of the loose rock block. 
After the determination of the geometry of the loose, separated rock block 
the stability of the underground opening is calculated by using beam 
theory as described hereinabove. Because roof deterioration is time 
dependent and gradual, repeated vacuum hole tests can be used to provide 
an early warning system to the mining engineer concerning an impending 
unstable roof condition. The vacuum method of the present invention is 
particularly suitable to monitor the deterioration of the fractured rock, 
because the air flow rate through a fracture increases by the cube of the 
width of the fracture. Unchanged pneumatic parameters indicate unchanged 
fracturation condition and vice versa. The formula to estimate the 
relative widening of the fracture by flow rate measurement using the 
original fracture width as unit is: 
##EQU6## 
where: F=fracture size ratio--(dimensionless) 
W.sub.o =original fracture size (mm) 
W.sub.i =increased fracture size (mm) 
Q.sub.o =original flow rate at differential pressure P(cc/m) 
Q.sub.i =increased flow rate at differential pressure P(cc/m). 
It will, of course, be appreciated that when testing a mine roof for 
separation, application of a positive or hydrostatic pressure to 
observation hole 1 would impose an additional load on the rock which could 
be sufficient to cause roof collapse, and for this reason a vacuum or 
negative pressure is always applied to the observation hole when testing 
roofs. The invention is, however, equally applicable to the testing of 
side walls and the bottom or floor of the stope, and in these areas the 
application of a positive pressure does not impose such a safety hazard 
and under certain circumstances may be preferred. It will be appreciated 
that in certain applications, the integrity of roof, walls and floors is 
of considerable importance. For example, selection of a site for 
underground disposal of radioactive waste material, imposes a need to 
establish the sealing ability of the rock against migration of 
radionuclides in all directions. If samples of the gas flowing to the 
vacuum hole are taken for either in-situ analysis or analysis at a 
selected remote laboratory, it is possible to determine the presence or 
absence of possibly dangerous gases in the rock mass and/or radioactive 
material in the mass. Similarly fracturation problems in connection with 
concrete objects, such as foundations or dams may be investigated using 
either positive or negative pressure in the observation hole. 
EXAMPLE 
In order to illustrate the present invention, on a laboratory scale, a 
block of unfractured uranium ore, 100.times.100.times.100 cm thick was 
tested and found to have a matrix permeability of 0.005 md. A 70 cm Hg 
vacuum was drawn on one side of the block, while the other side of the 
block was at atmospheric pressure, and the time measured for this vacuum 
to decay to zero, i.e. equalization of pressure on both sides of the 
block. It was found that pressure equalization occurred in 180 minutes. 
After completion of this portion of the experiment a circular hole 0.34 mm 
in diameter was drilled through the sample to simulate fracturation. The 
vacuum pressure test was again applied and it was found that the 70 cm Hg 
vacuum equalized to atmospheric pressure in 2 minutes, indicative of a 
permeability of 3,6000,000 md. It will be observed the time index, i.e. 
ratio of the equalization time in unfractured rock to the equalization 
time in fractured rock =180/2=90. It is obvious that such a drastic 
pneumatic condition change can be detected easily with very simple vacuum 
gauges. 
The present invention is also useful to evaluate the best way of locating 
and arranging rock bolt installation in a mine roof. The basic requirement 
to install a rock bolt properly is to anchor it in a competent section of 
the rock mass overlying the underground opening. In this case the rock 
bolt suspends effectively the fractured rock below. But heretofore no 
method has been available to check the proper placement of the rock bolt 
anchor. The procedure to check the placement of the rock bolt with the 
vacuum method of the present invention is: 
(1) determine the location of fractures present in the rock bolt holes 
prior to the installation of the rock bolt itself by using the method 
described hereinabove; 
(2) based on the location of the fractures establish the thickness of the 
continuous roof section which can be considered as a beam; (H) 
(3) calculate the thickness of the so-called self-supporting beam by using 
the formula: 
##EQU7## 
where: h=thickness of the self-supporting beam 
.phi.=allowable tensile stress in the beam 
.gamma.=unit weight of the rock 
S=span of the opening; 
(4) if the thickness of the continuous roof section (H) where the rock bolt 
anchor will be landed is larger than the thickness of the calculated 
self-supporting beam (h) the rock bolt installation is proper, i.e.: 
EQU H&gt;h, 
if the h&lt;H the rock bolt installation is not safe. 
As noted hereinabove, fracture parameter determination in some cases 
requires measurement of flow rates which are far below the measuring range 
of commercially available rotameters or the like. It has, therefore, been 
necessary to develop a special flow meter capable of directly measuring 
air flow rates as low as 0.001 cc per minute or less. As shown in FIG. 7, 
a suitable flow meter 16 comprises a vertical frame 71 having a tube 
holder 72 pivotally mounted on a horizontal axis 73 thereon. A calibrated 
glass measuring tube 74 is mounted on holder 72 and connected to an air 
flow via plastic tubing 75 and normally open valve 76. Before making a 
measurement, valve 76 is closed to the formation and opened to atmsophere 
via valve 69 and a drop of water 77 is introduced, by a syringe, into tube 
74 and positioned just beyond the zero mark by tilting tube 74 out of the 
horizontal. After placement the tube is restored to the horizontal 
position and valve 79 is closed while valve 76 is again opened. The flow 
rate is measured by recording the movement of the water drop along the 
tube, i.e., measuring the increase in volume during unit time. By using a 
water drop in a horizontal tube, rather than an air bubble in a vertical 
tube as in the well known bubblemeter, permits very low flow rates to be 
measured over measuring periods of the order of several days. At the low 
flow rates contemplated by the present invention, the specific weight of 
the flowing air can have a significant effect upon readings from a 
variable area type flow meter, and for this reason it is desirable to 
filter the exhaust air from the vacuum pump 9 (FIG. 5) before passing to 
horizontal flow meter 16, in order to eliminate oil mist from the air, 
through an exhaust assembly 17.