U.S. patent number 3,735,704 [Application Number 05/014,166] was granted by the patent office on 1973-05-29 for control blasting.
Invention is credited to Clifton W. Livingston.
United States Patent |
3,735,704 |
Livingston |
May 29, 1973 |
CONTROL BLASTING
Abstract
Disclosed herein is a new concept in blasting which is based on
a unique theory that permits total utilization of the energy of a
blasting charge. The method of this invention enables one, through
a set procedure, to control the direction and distance of placement
of huge volumes of earth and/or the degree of breakage of a volume
of material adjacent the area effected by the blast. The method
requires systematic test blasts to determine parameters for use in
established formula so that the desired end of control blasting of
huge volumes of earth can be attained. The scope of this invention
is unlimited in its application to mining, hydropower development,
construction projects, etc.
Inventors: |
Livingston; Clifton W. (Grand
Junction, CO) |
Family
ID: |
21763898 |
Appl.
No.: |
05/014,166 |
Filed: |
February 25, 1970 |
Current U.S.
Class: |
102/301 |
Current CPC
Class: |
E21C
41/31 (20130101); F42D 1/00 (20130101); E21C
37/16 (20130101); E21C 41/32 (20130101) |
Current International
Class: |
E21C
41/26 (20060101); F42D 1/00 (20060101); E21C
41/00 (20060101); F42d 001/00 () |
Field of
Search: |
;102/22,23 |
Foreign Patent Documents
|
|
|
|
|
|
|
1,200,232 |
|
Sep 1965 |
|
DT |
|
1,385,793 |
|
Jul 1964 |
|
FR |
|
Other References
Alan Bauer et al., How IOC Puts Crater Research to Work, E &
MJ, September, 1965. .
L. C. Lang, Mining Itabirite on the Bong Range in Liberia, Reprint,
Canadian Mining Journal, November, 1965. .
L. C. Lang, Pit Slope Control by Controlled Blasting, Reprint,
Canadian Mining Journal, December, 1965. .
L. C. Lang, Blasting Frozen Iron Ore at Knob Lake Reprint, Canadian
Mining Journal, August 1966..
|
Primary Examiner: Pendegrass; Verlin R.
Claims
What is claimed is:
1. In bench geometry blasting, where for the particular material to
be blasted and explosive to be employed, the values of d, d.sub.o,
N, .DELTA.o and E are known for a given weight of explosive charge
W,
d being any depth of burden of the charge;
d.sub.o being the depth of placement of a charge of the given
weight which produces upon explosion of the charge maximum pressure
rise within the material;
N being the depth of placement of a charge of the given weight
which first begins to produce upon explosion of the charge
fractures on the surface of the material in which the charge is
placed (as illustrated in accompanying FIG. 1);
.DELTA..sub.o being the value of d.sub.o /N for the given charge
weight; and
E being the value calculated from the formula N = E .cuberoot.W
using the value of N for the given weight;
and where for the same particular material to be blasted and
explosive to be employed, the values of d.sub.bo, N, and
.DELTA..sub.bo are known,
d.sub.bo being the perpendicular distance from the vertical free
face to the center of gravity of explosive charge which produces
upon explosion of said charge maximum pressure rise within the
material;
N being as previously defined for the same charge weight used in
determining d.sub.bo ; and
.DELTA..sub.bo being the value of d .sub.bo /N for the given charge
weight;
the improvement comprising controlling the trajectory of material
formed as bench geometry surrounding a blast charge in response to
explosion of the charge by;
a. placing a charge of weight W at a vertical depth .DELTA..sub.v
wherein .DELTA..sub.v = d.sub.v /N wherein d.sub.v is the vertical
depth;
b. positioning said charge of weight W along a rearward extension
of a line perpendicular to the bench free face at a burden
.DELTA..sub.b wherein .DELTA..sub.b = d.sub.b /N wherein d.sub.b is
the burden distance and .DELTA..sub.b = .DELTA..sub.v +
.DELTA..sub.diff wherein .DELTA..sub.diff is a predetermined
positive value, the trajectory being substantially in a vertical
direction when .DELTA..sub.v is less than .DELTA..sub.b and
substantially in the horizontal face direction when .DELTA..sub.b
is less than .DELTA..sub.v ; and
c. detonating the charge.
2. The method as defined in Claim 1, wherein the charge is placed
at a vertical depth beneath the surface equal to d.sub.b +
d.sub.diff d.sub.diff being calculated from a .DELTA..sub.diff
equalling d.sub.diff/ N and .DELTA..sub.diff is within the range of
0.02 - 0.20.
3. The method as defined in claim 1 wherein the known values and
.DELTA..sub.diff are determined by conducting test crater and bench
blasts.
4. In bench geometry blasting the method of controlling the
trajectory of material surrounding a blast charge in response to
explosion of the charge comprising:
a. Conducting test crater blasts to determine values for d.sub.o,
N, and E for the surrounding material and type of explosive, where
d.sub.o is the depth of placement of a charge of known weight which
produces upon explosion of the charge maximum pressure rise in the
material; where N is the depth of placement of a charge of a known
weight which first begins to produce upon explosion of the charge a
crater on the surface of the material in which the charge is placed
and where E is calculated from the formula N = E .cuberoot. W using
the determined value of N and its known charge weight;
b. Conducting test bench blasts to determine the value of
.DELTA..sub.bo for the surrounding material and type of explosive,
where .DELTA..sub.bo is d.sub.bo /N and d.sub.bo is the
perpendicular distance from the free face of a bench geometry to
the center of gravity of an explosive charge placed at a depth
d.sub.o which produces maximum pressure rise within the
material;
c. Placing a charge of weight determined by the formula d.sub.o =
.DELTA..sub.o E .cuberoot. W at a depth of d.sub.o + d.sub.diff
beneath the upper surface of the material bench, d.sub.diff being a
predetermined positive value;
d. Placing the charge along a rearward extension of the line of
preferred trajectory a distance equal d.sub.bo, whereby trajectory
is controlled to be substantially in the horizontal direction;
and
e. Detonating the charge.
5. The method as defined in claim 4, wherein d.sub.diff is based on
a .DELTA..sub.diff in the range of 0.02- 0.20 and .DELTA..sub.diff
= d.sub.diff /N.
6. A method according to claim 4 wherein d.sub.bo is determined by
placing a series of test blasts of known weight at depths d.sub.o
and at various distances d.sub.b from the the bench free face,
detonating said charges and measuring the response of the material
surrounding the blast to determine which burden distance produces
the maximum doming without venting in the direction of the bench
free face.
7. A method according to claim 4 wherein d.sub.bo is determined by
placing a series of test blasts of known weight at depths d.sub.o
and at various distances d.sub.b from the bench free face,
detonating said charges and measuring the response of the material
surrounding the blast to determine which burden distance produces
the maximum volume of material broken per pound of explosive.
8. In bench blasting, the method of controlling the trajectory of
the material surrounding the blast charge in response to explosion
of the charge including preparing the bench face to have a
disposition of substantially 45.degree. with the horizontal,
placing the charge along a line projecting perpendicular from
substantially mid-point of the free face but within a range on
either side of the free face equal to 70 percent of the free face
length a distance from the free face substantially equal to the
d.sub.bo for the material and explosive, d.sub.bo being the
perpendicular distance from the vertical free face to the center of
gravity of explosive charge which produces upon explosion of said
charge maximum pressure rise within the material, spacing the
charge a distance from the upper horizontal surface of the bench a
distance equal to d, wherein d is any depth of the charge when d is
determined from .DELTA..sub.o + .DELTA..sub.diff + .DELTA..sub.inc
and .DELTA. is related to d by the formula .DELTA. = d/N,
.DELTA..sub.o being the value of d.sub.o /N for the given charge
weight, .DELTA..sub.diff being a predetermined positive value,
.DELTA..sub.inc being .DELTA..sub.bo minus .DELTA..sub.o,
.DELTA..sub.bo being the value of d.sub.bo /N for the given charge
weight, d.sub.o being the depth of placement of a charge of the
given weight which produces upon explosion of the charge maximum
pressure rise within the material, N being the critical depth (as
described in conjunction with FIG. 2), .DELTA..sub.diff being in
the range of 0.05 to 0.25.
9. The method according to claim 8, wherein the distance from the
free face is in the range of d.sub.bo to d.sub.b where d.sub.bo is
a function of .DELTA..sub.bo and d.sub.b is a function of
.DELTA..sub.bo - 0.15 and .DELTA..sub.bo is related by d.sub.bo by
the formula .DELTA..sub.bo = d.sub.bo /N.
10. The method of autogenous blasting to produce a fine grain
particle comprising:
a. forming a crater by detonating a first explosive charge, the
explosive set beneath the surface of earthen material to produce a
muck pile backed against the free face formed by the blast;
b. detonating at least one second explosive charge adjacent the
first named charge, the distance of the second charge from the free
face created by the first explosion being in the range where
implosion occurs and the vortex is well developed and is directed
toward the muck pile and corresponds to d values defined by
.DELTA..sub.bo to .DELTA..sub.a where .DELTA. = d/N and
.DELTA..sub.bo and .DELTA..sub.a are determined from test crater
blasts, .DELTA..sub.bo being the value of d.sub.bo /N for the given
charge weight, d.sub.bo being the perpendicular distance from the
vertical free face to the center of gravity of explosive charge
which produces upon explosion of said charge maximum pressure rise
within the material, N being the depth of placement of a charge of
the given weight which first begins to produce upon explosion of
the charge a crater on the surface of the material in which the
charge is place, .DELTA..sub.a being the burden distance at which
the velocity of the vortex is maximum.
11. The method as defined in claim 10 wherein the depth of the
charge is controlled so as to predetermine breakage limits of the
second blast whereby the angle of inclination .alpha. of a
subsequent blast is controlled within the range of .alpha.
o.degree. to 45.degree. .
12. The method of controlling the granulometry of material
surrounding a charge of explosive in response to explosion of the
charge comprising:
a. conducting test crater blasts at various depths to determine the
breakage limits of the material in response to the blasts;
b. preparing a chart by plotting the slant distance to each of the
types of fractures in the blasting failure process and to the
doming limit from test crater blasts;
c. establishing a grid over the crater formed by each test blast
and measuring the particle size of the broken material touched by
the grid;
d. preparing a chart by plotting the variation of the predominate
particle size with various charge depths at scaled depth ratios
.DELTA.;
e. determining in the range .DELTA..sub.d to .DELTA..sub.o from the
chart of (d) the value in the controlling burden direction at which
the desired predominate particle size can be achieved,
.DELTA..sub.d is the .DELTA. value at which doming begins and
.DELTA..sub.o is the value of d.sub.o /N for the given charge
weight, d.sub.o being the depth of placement of a charge of the
given weight which produces upon explosion of the charge maximum
pressure rise within the material, and N being the depth of
placement of a charge of the given weight which first begins to
produce upon explosion of the charge fractures on the surface of
the material in which the charge is placed; or in the range
.DELTA..sub.o to .DELTA. equals 0 the value at which the desired
particle size can be achieved where predominant particle size from
autogenous blasting ranges from near zero to that measured at
.DELTA. .sub.o of the crater tests;
f. decreasing the crater chart value (e) by subtracting for bench
geometry a .DELTA. .sub.increment adjustment ranging linearly from
a maximum at .DELTA..sub.bo -.DELTA..sub.o for a controlling burden
of magnitude .DELTA..sub.o to zero for a controlling burden
.DELTA..sub.a or less or for a controlling burden .DELTA..sub.ss or
greater, .DELTA. .sub.increment being .DELTA. .sub.bo minus .DELTA.
.sub.o, .DELTA..sub.bo being the value of d.sub.bo /N for the given
charge weight, d.sub.bo being the perpendicular distance from the
vertical free face to the center of gravity of explosive charge
which produces upon explosion of said charge maximum pressure rise
within the material, .DELTA..sub.a is the .DELTA. value at which
the volume of the apparent crater is maximum, and .DELTA..sub.ss
being the .DELTA. value at which cratering begins;
g. placing the charge at a distance from the controlling burden
toward the face of the bench corresponding to the adjusted .DELTA.
value;
h. placing the charge at a distance from the non-controlling free
face at the crater test determined value; and
i. detonating the charge.
13. The method of stability control bench blasting, including:
a. conducting test crater blasts and preparing a chart by plotting
the slant distance to limits of each of the types of fractures in
the blasting failure process and to the doming limit, by plotting
the slant distances to the crater lip against charge depth and the
zone of bond strength destruction against charge depth, R cracks
defining the limits of extreme rupture and by plotting the
variation in predominant particle size against charge depth from
the parameters determined from said blasts;
b. selecting a predetermined limit part which the material in which
the blasts are conducted is not to be affected;
c. progressively setting and detonating explosive charges in the
direction approaching said limit, each successive charge being
decreased in .DELTA..sub.b value and increased in .DELTA..sub.v
value so as to progressively increase the .DELTA..sub.v
-.DELTA..sub.b difference approaching said limit, the .DELTA..sub.b
value being determined from the cratering results using the
relation b/N, where b is the bench burden in feet and N is the
distance in feet at which cratering begins for the bench charge,
the .DELTA..sub.v value being determined from the relation d.sub.c
/N, where d.sub.c is the distance beneath the top of bench in feet
to the bench charge; and
d. the placement of the charge being such that the limits of
extreme rupture as defined in said chart do not for any given blast
extend beyond the predetermined limit.
14. The method as defined in claim 13, wherein the burden of any
charge approaching said limit is of a distance d.sub.b controlled
by .DELTA., and .DELTA. is in the range .DELTA..sub.0 - 0.10 and
.DELTA..sub.0 -0.20, and where .DELTA. =d.sub.b /N.
15. The method as defined in claim 13, further including the step
of placing rock bolts in the material prior to blasting to
reinforce the material beyond said limit.
16. The method as defined in claim 13, further including digging
breakage limit trenches in the surface of the surrounding material
along the line of the predetermined limit.
17. The method as defined in claim 1 wherein the charge is placed
at a controlling burden from the vertical free face equal to
d.sub.v +d .sub.diff being calculated from a .DELTA..sub.diff
equalling d.sub.diff /N and .DELTA..sub.diff is within the range of
0.20 - 0.20.
18. The method as defined in claim 3 wherein said test crater and
bench blasts include:
a. conducting cratering tests so as to measure the limits of R
cracks, I cracks, O cracks, shock slabs, rebound slabs, crater and
doming;
b. measuring time, velocity, and duration of massive ground motion
for each of the charges of said cratering tests and determining the
variation with scaled charge depth for the series of charges for a
given type, shape, weight and depth of charge;
c. excavating the craters formed in said cratering tests so as to
measure the limits of the bond strength destruction zone,
intermediate zone, and the barrier zone and to determine the
granulometry, size and shape of particles in each zone and for each
blast, and summarizing the variation for the series of blasts with
scaled charge depth;
d. conducting test bench blasts so as to:
i. observe the effect of the second free face upon breakage limits,
particle size, and volume of material broken per pound of
explosive;
ii. determine the variation in time, velocity and duration of
massive ground motion in the .DELTA..sub.v direction towards the
top of the bench and in the .DELTA..sub.b direction towards the
face of the bench at various values of .DELTA..sub.v -
.DELTA..sub.b, and
iii. determine the scaled slant distance to the doming limit at
various values of .DELTA..sub.b, the controlling burden;
e. determine the .DELTA..sub.increment adjustment from said test
bench blasts and applying the adjustment to correct fracture,
crater, doming limits, particle size, granulometry and shape to the
geometry of bench blasting;
f. utilizing a chart plotting and slant distances to the doming
limit for a given blast to predetermine excavation limits in bench
blasting to determine .DELTA..sub.v given .DELTA..sub.b, or to
determine .DELTA..sub.b given .DELTA..sub.b so as to predetermine
the position and inclination of the free face prior to a bench
blast or successive bench blasts;
g. predetermining the time, velocity, duration and trajectory of a
blast in the .DELTA..sub.b direction towards the face of the bench
and the .DELTA..sub.v direction towards the top of the bench so as
to achieve the desired throw in either or both directions, such
determination of the throw being based upon the results of the
crater tests; and
h. predetermining excavation limits as in (a) and simultaneously
controlling trajectory and granulometry towards the top and towards
the face of the bench so as to predetermine both trajectory and
granulometery from results of the crater and bench tests.
19. A method of trajectory control blasting comprising:
a. conducting crater and bench blast tests to predetermine limits
of fracturing, cratering and doming in bench blasting and so as to
determine the variation in time, velocity and duration of massive
ground motion with variation in controlled burden and variation in
.DELTA. difference, .DELTA..sub.v -.DELTA..sub.b, wherein
.DELTA..sub.v = d.sub.v /N wherein d.sub.v is the vertical depth
and N is the depth of placement of a charge of the given weight
which first begins to produce upon explosion of the charge
fractures on the surface of the material in which the charge is
placed, and .DELTA..sub.b = d.sub.b /N wherein d.sub.b is the
burden distance;
b. creating a free face at a predetermined angle perpendicular to
the angle of desired trajectory of a subsequent blast by placing
and detonating a charge of known weight at a burden, .DELTA..sub.
b, and depth of center of gravity of the explosive charge,
.DELTA..sub.v, at which the inclination of a line from the center
of gravity of the charge to the doming limit of the bench test
produces the desired inclination;
c. determining .DELTA..sub.v, .DELTA..sub.b, and .DELTA..sub.v
-.DELTA..sub.b from cratering and bench tests and placing the
charge so that:
i. the desired proportions of massive ground motion both in the
.DELTA..sub.v direction towards the top of the bench and in the
.DELTA..sub.b direction towards the face of the bench are
predetermined to produce the desired trajectory, granulometry, or
both, and
ii. so that the time, velocity and duration of massive ground
motion will be such as to produce the desired throw;
d. placing the charge on the rearward extension of the line of
desired trajectory at a distance .DELTA..sub.b so as to obtain the
desired breakway velocity in a direction normal to the free face
and as indicated by the bench tests, and so as to predetermine the
time, velocity and duration of massive ground motion in the
.DELTA..sub.v direction toward the top of the bench; and
e. detonating the charge.
20. A method of maximizing the quantity of fine material from a
blast including:
a. conducting crater blast tests so as to determine limits of R
cracks, I cracks, O cracks, slabs, rebound slabs, and limits both
of the apparent crater and the true crater, and to determine the
variation in predominant particle size with scaled charge depth and
the time, velocity and duration of massive ground motion, including
excavating the craters to observe the limits of bond strength
destruction, the intermediate zone, and the barrier zone, and
determining by observation of cratering results the depth at which
the explosion cavity collapses and a ground motion vortex action
begins;
b. preparing a chart by plotting the slant distances to limits of
each of the types of fractures in the blasting failure process and
to the doming limit, by plotting the slant distances to the crater
lip against charge depth and the zone of bond strength destruction
against charge depth, R cracks defining the limits of extreme
rupture and by plotting the variation in predominant particle size
against charge depth from the parameters determined from said
blasts;
c. preparing an opening cut or free face by blasting so as to
predetermine the inclination of the sides within the range of
45.degree. from the horizontal to vertical and so as to produce a
muckpile to inhibit ground motion of subsequent blasts;
d. placing a charge of given type of explosive of given weight, W,
behind the face so that the distance .DELTA..sub.v beneath the top
of the bench is in the range of .DELTA..sub.o -0.02 to
.DELTA..sub.o -0.10 and the distance .DELTA..sub.b perpendicular to
the face is in the range .DELTA.= .DELTA..sub.mb to (.DELTA. .sub.a
+ .DELTA..sub.o)/2 where .DELTA..sub.a is the charge depth at which
the volume of the apparent crater is maximum, .DELTA..sub.o is the
charge depth at which the volume of the true crater is maximum,
.DELTA..sub.mb is the charge depth at which the slant distance from
the charge to the limit of bond strength destruction is maximum;
the weight of the charge W, the depth of charge d.sub.c, and the
burden b being determined by the equations
i. N = E .cuberoot. W wherein N is the depth of placement of a
charge of the given weight which first begins to produce upon
explosion of the charge fractures on the surface of the material in
which the charge is placed, and E is the value calculated using the
value of N for the given weight,
ii. .DELTA..sub.b = b/N,and .DELTA..sub.v = d.sub.c /N,
iii. d.sub.c = .DELTA.E .cuberoot. W, and
iv. V/W = E.sup.3 ABC, wherein A = V/V.sub.o, V being the crater
volume, V.sub.o being the crater volume at optimum charge depth d,
B = V.sub.o /N.sup.3, and C = (V/V.sub.ob).sup.. C.sub.o wherein
C.sub.o = (V/W bench)/(V.sub.o /W crater);
e. placing said charge such that a vortex action is produced as a
result of implosion and ground motion following collapse of the
explosion cavity, and the venting gas bubble is directed into the
muckpile so as to develop a scouring action between particles and
so as to minimize excessive flyrock which otherwise would occur,
said vortex action occurring within the range .DELTA. = 0 to
.DELTA. = .DELTA..sub.o, the venting gas bubble being controlled in
the direction of .DELTA..sub.b, if .DELTA..sub.v is sufficiently
larger than .DELTA..sub.b, and the scouring action between
particles is developed for the particular controlling value of
.DELTA..sub.b if the duration of the action equals that measured
for a cratering blast of the same explosive at the same .DELTA.
value and charge shape; and
f. detonating the charge.
21. The method as defined in claim 20 including
a. directing the vortex action into the muckpile by assuring that
ground motion towards the opening cut or inclined free face will
predominate over that in the vertical direction by predetermining
.DELTA..sub.v to exceed .DELTA..sub.b and defining the
.DELTA..sub.b -.DELTA..sub.v difference to be within the
limitations: .DELTA..sub.b is greater than .DELTA..sub.mb and less
than (.DELTA..sub.a + .DELTA..sub.o)/2 and .DELTA..sub.v is greater
than .DELTA..sub.o -0.02 and less than or equal to .DELTA..sub.0
-0.10;
b. predetermining the particle size attributable to ground motion
in the .DELTA..sub.b direction to be smaller than that attributable
to ground motion in the .DELTA..sub.v direction due to:
i. the scouring action within the vortex attributable to the
greater velocity near the center than the edges of the vortex,
and
ii. containment of the vortex due to the .DELTA..sub.v
-.DELTA..sub.b difference; and
c. predetermining the particle size in the .DELTA..sub.b and
.DELTA..sub.v directions using a chart plotting predominant
particle size, zones and charge depth from test crater blasts.
22. The method as defined in claim 20 including:
a. determining the spacing between rows of holes of a multiple hole
blast by: controlling the position and inclination of the free
face; predetermining the excavation limit from the preceding row of
holes using results of the cratering tests and plotting a chart of
the variation in slant distance to crater, cracks and doming limits
against charge depth; and placing the charge behind the free face
next to be created so that .DELTA..sub.b is greater than
.DELTA..sub.mb and .DELTA..sub.b is less than (.DELTA..sub.a +
.DELTA..sub.o)/2 and .DELTA..sub.v is greater than .DELTA..sub.o
-0.02 and less than or equal to .DELTA..sub.o -0.10;
b. minimizing random ground motion by controlling the delay
interval between successive rows of holes so that the velocity of
massive ground motion in the vertical direction will not exceed
that as determined from cratering blast tests at a charge depth for
the range of .DELTA..sub.v previously specified;
c. minimizing random ground motion by detonating charges of
successive rows in such a sequence that a barrier mass determined
by the .DELTA..sub.diff of thickness not less than 1.4N exists or
that the barrier mass is destroyed by detonating charges in it at
the same instant as the row blast holes approaching the barrier
mass reaches a distance of 1.4N or less;
d. promoting the vortex action between successive rows of holes so
that the delay interval between successive rows will not be less
than the duration of ground motion as determined by a crater blast
test for a charge of the same shape, type and weight of explosive
at the same controlling burden;
e. minimizing excessive packing of the muckpile by varying the
angle of inclination .alpha. of the vortex within the range 0 to
45.degree. from the horizontal by predetermining the inclination of
successive free faces by predetermining with the controlling burden
the slant distance from the charge to the doming limit at the top
of the bench; and
f. increasing the proportion of fine material available to the
vortex action by controlling the spacing .DELTA..sub.s between
holes of a given row within the range .DELTA..sub.s = .DELTA..sub.b
to .DELTA..sub.s = 1.4.DELTA. .sub.b where .DELTA..sub.b previously
has been defined as being in the range .DELTA..sub.mb to
(.DELTA..sub.a + .DELTA..sub.o)/2, which are approximate limits of
the bond strength destruction zone as determined from the test
crater blasts.
23. The method as defined in claim 12 wherein said test crater and
bench blasts are conducted to:
a. determine parameters A and B using test crater blasts wherein A
= V/V.sub.o wherein V = the crater volume and V.sub.0 = the crater
volume at optimum charge depth d, and B = V.sub.o /N.sup. 3 ;
b. determine various values of parameter C using test bench blasts
for various heights of bench, burden and .DELTA..sub.v -
.DELTA..sub.b difference, wherein .DELTA..sub.v = d.sub.v /N
wherein d.sub.v is the vertical depth and .DELTA..sub.b = d.sub.b
/N wherein d.sub.b is the burden distance, and C = (V/V.sub.ob)
.sup.. C.sub.o wherein C.sub.o = (V/W bench)/(V.sub.o /W crater)
wherein V.sub.ob is the optimum crater volume in the burden
direction and the .DELTA. and V.sub.o are the volume of the crater
in the depth direction and burden direction, respectively; and
c. determine the geometry of the bench for a given explosive,
weight and shape of charge from measurements of breakage limits and
particle size in bench blasting and identified by the product
control index number, ABC of the equation V/W = E.sup.3 ABC, where
E is calculated from the formula N = E .cuberoot.W.
24. A method of multiple row product control blasting consisting
of:
a. conducting crater and bench blast tests to predetermine limits
of fracturing, cratering and doming in bench blasting and so as to
determine the variation in time, velocity and duration of massive
ground motion with variation in controlling burden and variation in
.DELTA..sub.difference, .DELTA..sub.v -.DELTA..sub.b, wherein
.DELTA..sub.v = d.sub.v /N wherein d.sub.v is the vertical depth
and N is the depth of placement of a charge of the given weight
which first begins to produce upon explosion of the charge
fractures on the surface of the material in which the charge is
placed, and .DELTA..sub.b = d.sub.b /N wherein d.sub.b is the
burden distance;
b. creating successive free faces at a predetermined angle by
placing a charge at a depth .DELTA..sub.v and a burden
.DELTA..sub.b and at a .DELTA..sub.diff, .DELTA..sub.diff being a
predetermined positive value, the trajectory being substantially in
the vertical direction when the burden equals .DELTA..sub.v +
.DELTA..sub.diff and substantially in the horizontal direction when
the depth equals .DELTA..sub.b + .DELTA..sub.diff, thereby allowing
predetermining of the controlling burden for the succeeding rows of
holes;
c. predetermining from results of the crater and bench tests from
the proper delay interval between blasts using time, velocity and
duration of ground motion measurements so that motion of a given
row does not begin before motion of the previous row is
complete;
d. placing a charge or series of charges with respect to the top of
the bench and the face of the bench so as to predetermine for each
charge the proportions of massive ground motion in each of the
depth and burden directions according to the granulometry desired;
and
e. successively detonating charges of the series at the
predetermined delay interval between charges.
25. A method of stability control blasting including:
a. conducting crater and bench blast tests to predetermine limits
of fracturing, cratering and doming in bench blasting and so as to
determine the variation in time, velocity and duration of massive
ground motion with variation in controlling burden and variation in
.DELTA..sub.difference, .DELTA..sub.v -.DELTA..sub.b, wherein
.DELTA..sub.v = d.sub.v /N wherein d.sub.v is the vertical depth
and N is the depth of placement of a charge of the given weight
which first begins to produce upon explosion of the charge
fractures on the surface of the material in which the charge is
placed, and .DELTA..sub.b = d.sub.b /N wherein d.sub.b is the
burden distance;
b. defining a predetermined limiting part which the material in
which the blasts are conducted is not to be affected, and defining
a stability control zone between the limiting part in the main body
to be excavated such that breakage limits and doming due to
blasting within said stability control zone will not extend beyond
limits of the stability control zone;
c. progressively setting and detonating explosive charges within
the stability control zone approaching the limiting part so that
the .DELTA..sub.v -.DELTA..sub.b difference progressively increases
so as to progressively suppress massive ground motion towards the
top of the bench and so as to progressively increase the velocity
of ground motion towards the face of the bench, .DELTA..sub.v =
d.sub.v /N wherein d.sub.v is the vertical depth and N is the depth
of placement of a charge of the given weight which first begins to
produce upon explosion of the charge a crater on the surface of the
material in which the charge is placed, and .DELTA..sub.b = d.sub.b
/N wherein d.sub.b is the burden distance; and
d. placing the charge such that .DELTA..sub.v exceeds the slant
distance to R cracks or to the doming limit, and rebound cracks do
not for any given blast extend beyond the predetermined limits.
26. A method of stability control blasting including:
a. conducting bench blast tests to predetermine limits of
fracturing, cratering and doming in bench blasting and so as to
determine the variation in time, velocity and duration of massive
ground motion with variation in controlling burden and variation in
.DELTA..sub.difference, .DELTA..sub.v -.DELTA..sub.b, wherein
.DELTA..sub.v = d.sub.v /N wherein d.sub.v is the vertical depth
and N is the depth of placement of a charge of the given weight
which first begins to produce upon explosion of the charge
fractures on the surface of the material in which the charge is
placed, and .DELTA..sub.b = d.sub.b /N wherein d.sub.b is the
burden distance; and to determine the .DELTA..sub.v -.DELTA..sub.b
difference at which massive ground motion towards the top of the
bench is insufficient to cause blasting damage at the collar of the
borehole in row blasting;
b. measuring the time, velocity and duration of massive ground
motion in the .DELTA..sub.b direction at the .DELTA..sub.v
-.DELTA..sub.b difference, as defined in step (a);
c. progressively setting and detonating charges in the direction of
mining advance based upon time, velocity and duration measurements
of the bench blast tests so that ground motion toward the top of
the bench is not predetermined by the duration of horizontal ground
motion of the previous charge; and
d. placing the charge such that .DELTA..sub.b equals the burden
determined experimentally for the .DELTA..sub.v -.DELTA..sub.b
difference, determined in step (a), and the delay interval between
successive charges is such that horizontal ground motion from
successive charges does not affect vertical ground motion, and the
weight of explosive detonated per delay does not exceed that for
which said .DELTA..sub.v -.DELTA..sub.b was determined.
27. The method as defined in claim 25 wherein the .DELTA..sub.v
-.DELTA..sub.b difference is determined within the range
.DELTA..sub.v is greater than 1.0 and less than 1.4, and
.DELTA..sub.b is within the range .DELTA..sub.o -0.10 and
.DELTA..sub.o -0.20.
28. In bench geometry blasting, the method of controlling the
trajectory of material surrounding a blast charge in response to
explosion of the charge comprising:
a. Conducting test crater blasts to determine values for d.sub.o,
N, and E for the surrounding material and type of explosive, where
d.sub.o is the depth of placement of a charge of known weight which
produces upon explosion of the charge maximum pressure rise in the
material; where N is the depth of placement of a charge of a known
weight which first begins to produce upon explosion of the charge a
crater on the surface of the material in which the charge is
placed; and where E is calculated from the formula N = E
.cuberoot.W using the determined value of N and its known charge
weight, said test crater blasts being carried out by setting
charges of known weight, shape and type of explosive at various
depths, detonating charges, measuring the limits of R cracks, I
cracks, O cracks, shock slabs, rebound slabs, cratering and doming
on the surface; and excavating the crater so as to determine the
shape of the true crater, limits of bond strength destruction zone,
the intermediate zone, barrier zone and transition limits;
b. Conducting test bench blasts to determine the value of
.DELTA..sub.bo for the surrounding material and type of explosive,
where .DELTA..sub.bo is d.sub.bo /N and d.sub.bo is the
perpendicular distance from the free face of a bench geometry to
the center of gravity of an explosive charge placed at a depth
d.sub.o which produces maximum pressure rise within the
material;
c. Placing a charge of weight determined by the formula d.sub.o =
.DELTA..sub.o E .cuberoot.W at a depth d.sub.o + d.sub.diff beneath
the upper surface of the material bench, d.sub.diff being a
predetermined positive value;
d. Placing the charge along a rearward extension of the line of
preferred trajectory a distance equal d.sub.bo, whereby trajectory
is controlled to be substantially in the horizontal direction;
and
e. Detonating the charge.
29. In bench geometry blasting, the method of controlling the
trajectory of material surrounding a blast charge in response to
explosion of the charge comprising:
a. Conducting test crater blasts to determine values for d.sub.o,
N, and E for the surrounding material and type of explosive, where
d.sub.o is the depth of placement of a charge of known weight which
produces upon explosion of the charge maximum pressure rise in the
material; where N is the depth of placement of a charge of a known
weight which first begins to produce upon explosion of the charge a
crater on the surface of the material in which the charge is placed
and where E is calculated from the formula N = E .cuberoot.W using
the determined value of N and its known charge weight, said test
crater blasts being performed by detonating a series of charges of
given type and known shape and weight at various depths measuring
the ground rise to determine the time, velocity and duration of
ground motion to differentiate between effects attributable to the
arrival of the shock wave, the pressure rise in the material, and
the venting of the gas bubble;
b. Conducting test bench blasts to determine the value of
.DELTA..sub.bo for the surrounding material and type of explosive,
where .DELTA..sub.bo is (d.sub.bo /N) and d.sub.bo is the
perpendicular distance from the free face of a bench geometry to
the center of gravity of an explosive charge placed at a depth
d.sub.o which produces maximum pressure rise within the
material;
c. Placing a charge of weight determined by the formula d.sub.o =
.DELTA. .sub.o E .cuberoot.W at a depth d.sub.o + d.sub.diff
beneath the upper surface of the material bench, d.sub.diff being a
predetermined positive value;
d. Placing the charge along a rearward extension of the line of
preferred trajectory a distance equal d.sub.bo, whereby trajectory
is controlled to be substantially in the horizontal direction;
and
e. Detonating the charge.
30. In bench geometry blasting, the method of controlling the
trajectory of material surrounding a blast charge in response to
explosion of the charge comprising:
a. Conducting test crater blasts to determine values for d.sub.o,
N, and E for the surrounding material and type of explosive, where
d.sub.o is the depth of placement of a charge of known weight which
produces upon explosion of the charge maximum pressure rise in the
material; where N is the depth of placement of a charge of a known
weight which first begins to produce upon explosion of the charge a
crater on the surface of the material in which the charge is
placed; and where E is calculated from the formula N = E
.cuberoot.W using the determined value of N and its known charge
weight;
b. Conducting test bench blasts to determine the value of
.DELTA..sub.bo for the surrounding material and type of explosive,
where .DELTA..sub.bo is (d.sub.bo /N) and d.sub.bo is the
perpendicular distance from the free face of a bench geometry to
the center of gravity of an explosive charge placed at a depth
d.sub.o which produces maximum pressure rise within the material,
said test crater and bench blasts being conducted to:
i. determine parameters A and B using test crater blasts wherein A
= (V/V.sub. o) wherein V = the crater volume and V.sub.o = the
crater volume at optimum charge depth d, and B = (V.sub.o
/N.sup.3);
ii. determine various values of parameter C using test bench blasts
for various heights of bench, burden and .DELTA..sub.v
-.DELTA..sub.b difference, wherein .DELTA..sub.v = (d.sub.v /N)
wherein d.sub.v is the vertical depth and .DELTA..sub.b = (d.sub.b
/N) wherein d.sub.b is the burden distance, and C = (V/V.sub. ob)
.sup.. C.sub.o wherein C.sub.o = (V/W bench)/(V.sub.o /W
crater);
iii. determine the geometry of the bench for a given explosive,
weight and shape of charge from measurements of breakage limits and
particle size in bench blasting and identified by the product
control index number, ABC of the equation (V/W) = E.sup.3 ABC,
where E is calculated from the formula N = E .cuberoot. W;
c. Placing a charge of weight determined by the formula d.sub.o =
.DELTA..sub.o E .cuberoot. W at a depth d.sub.o + d.sub.diff
beneath the upper surface of the material bench, d.sub.diff being a
predetermined positive value;
d. Placing the charge along a rearward extension of the line of
preferred trajectory a distance equal d.sub.bo, whereby trajectory
is controlled to be substantially in the horizontal direction;
and
e. Detonating the charge.
31. In bench geometry blasting, the method of controlling the
trajectory of material surrounding a blast charge in response to
explosion of the charge comprising:
a. Conducting test crater blasts to determine values for d.sub.o,
N, and E for the surrounding material and type of explosive, where
d.sub.o is the depth of placement of a charge of known weight which
produces upon explosion of the charge maximum pressure rise in the
material; where N is the depth of placement of a charge of a known
weight which first begins to produce upon explosion of the charge a
crater on the surface of the material in which the charge is
placed; and where E is calculated from the formula N = E .cuberoot.
W using the determined value of N and its known charge weight;
b. Conducting test bench blasts to determine the value of
.DELTA..sub.bo for the surrounding material and type of explosive,
where .DELTA..sub.bo is (d.sub.bo /N) and d.sub.bo is the
perpendicular distance from the free face of a bench geometry to
the center of gravity of an explosive charge placed at a depth
d.sub.o which produces maximum pressure rise within the
material;
c. Conducting crater blasts so as to determine the breakaway
velocity at .DELTA..sub.o as to identify granulometry
representative of each of the bond strength destruction zone, the
intermediate zone and the barrier zone surrounding the charge;
d. Conducting bench blasts to determine .DELTA..sub.bo and
.DELTA..sub.increment adjustment .DELTA..sub.bo -.DELTA..sub.o, by
placing a charge of substantially spherical shape at depth,
.DELTA..sub.o, as determined by the crater blasting and varying the
burden in the range .DELTA..sub.b greater than .DELTA..sub.o to
determine the .DELTA..sub.increment, .DELTA..sub.bo -.DELTA..sub.o
where (V/W) is maximum and the breakaway velocity towards the face
of the bench equals that toward the top of the bench;
e. Correcting cratering granulometry, limits of the bond strength
destructive zone, the intermediate zone, the barrier zone and
limits of the crater to the geometry of bench blasting;
f. Placing a charge of weight determined by the formula d.sub.o =
.DELTA..sub.o E .cuberoot. W at a depth d.sub.o + d.sub.diff
beneath the upper surface of the material bench, d.sub.diff being a
predetermined positive value;
g. Placing the charge along a rearward extension of the line of
preferred trajectory a distance equal d.sub.bo, whereby trajectory
is controlled to be substantially in the horizontal direction;
and
h. Detonating the charge.
32. A method of stability control blasting including:
a. Conducting crater and bench blast tests to predetermine limits
of fracturing, cratering and doming in bench blasting and so as to
determine the variation in time, velocity and duration of massive
ground motion with variation in controlling burden and variation in
.DELTA..sub.difference, .DELTA..sub.v -.DELTA..sub.b, wherein
.DELTA..sub.v = (d.sub.v /N) wherein d.sub.v is the vertical depth
and N is the depth of placement of a charge of the given weight
which first begins to produce upon explosion of the charge
fractures on the surface of the material in which the charge is
placed, and .DELTA..sub.b = (d.sub.b /N) wherein d.sub.b is the
burden distance;
b. Defining a predetermined limiting part which the material in
which the blasts are conducted is not to be affected, and defining
a stability control zone between the limiting part in the main body
to be excavated such that breakage limits and doming due to
blasting within said stability control zone will not extend beyond
limits of the stability control zone;
c. Progressively setting and detonating explosive charges within
the stability control zone approaching the limiting part so that
the .DELTA..sub.v -.DELTA..sub.b difference progressively increases
so as to progressively suppress massive ground motion towards the
top of the bench and so as to progressively increase the velocity
of ground motion towards the face of the bench, .DELTA..sub.v =
(d.sub.v /N) wherein d.sub.v is the vertical depth and N is the
depth of placement of a charge of the given weight which first
begins to produce upon explosion of the charge a crater on the
surface of the material in which the charge is placed, and
.DELTA..sub.b = (d.sub.b /N) wherein d.sub.b is the burden
distance;
d. Placing the charge such that .DELTA..sub.v exceeds the slant
distance to R cracks or to the doming limit, and rebound cracks do
not for any given blast extend beyond the predetermined limits;
e. Excavating a breakage limit trench, one wall of which coincides
with said limiting part;
f. Placing rock bolts into the permanent wall along the line of the
limiting part; and
g. backfilling the breakage limit trench prior to production
blasting so that the fill material both acts as a shock dampening
substance and prevents fractures from developing beyond the
predetermined limit.
33. In bench geometry blasting, the method of controlling the
trajectory of material surrounding a blast charge in response to
explosion of the charge comprising:
a. Conducting test crater blasts to determine values for d.sub.o,
N, and E for the surrounding material and type of explosive, where
d.sub.o is the depth of placement of a charge of known weight which
produces upon explosion of the charge maximum pressure rise in the
material; where N is the depth of placement of a charge of a known
weight which first begins to produce upon explosion of the charge a
crater on the surface of the material in which the charge is
placed; and where E is calculated from the formula N = E .cuberoot.
W using the determined value of N and its known charge weight;
b. Conducting test bench blasts to determine the value of
.DELTA..sub.bo for the surrounding material and type of explosive,
where .DELTA..sub.bo is (d.sub.bo /N) and d.sub.bo is the
perpendicular distance from the free face of a bench geometry to
the center of gravity of an explosive charge placed at a depth
d.sub.o which produces maximum pressure rise within the
material;
c. Placing a charge of weight determined by the formula d.sub.o =
.DELTA..sub.o E .cuberoot. W at a depth d.sub.o beneath the upper
surface of the material bench;
d. Placing the charge along a rearward extension of the line where
trajectory is to be substantially inhibited a distance equal to
d.sub.bo + d.sub.diff, d.sub.diff being a predetermined positive
value, whereby trajectory is controlled to be substantially in the
vertical direction; and
e. Detonating the charge.
Description
BACKGROUND OF THE INVENTION
Classification
This invention relates to blasting and particularly a blasting
method that enables one to control the trajectory, granulometry and
stability of the material surrounding the charge and the use of
these concepts in new and novel mining and hydropower development
methods.
Blasting and Mining History
The art of mining has a history that extends back to the earliest
recorded history of man. The introduction of explosives to the art
of mining presented a major technological advance, but the use of
explosives in mining has to this point been dominated by the
absence of any scientific basis for their deployment. While it
would appear that from the time when the Chinese originally began
to experiment with explosives and with present day engineering
capabilities, the reaction to explosives would be known and
predictable, such is not the case; and the use of explosives in
mining has been significantly inhibited since their deployment is
based on a "hit-and-miss" approach and consequently the high cost
of explosives has limited their use in mining, primarily to
breakage applications.
Since the reaction to blasting and the basis therefor has not
heretofore been known, the evaluation of blasting applications and
the logical, systematic analysis of problems involving such
applications in mining and other methods has not been possible.
The lack of an understanding of the reaction to blasting has made
it impossible to effectively utilize the energy of the explosive.
The new technology which I have evolved permits scientific analysis
of the problems of blasting, and I have developed a trajectory
control blasting method based on these principles.
Further, it has been generally considered that the breakage by
blasting of the materials surrounding the charge was a function of
the characteristics of the material, and that the lines of breakage
are dictated by the lines of weakness in the surrounding material.
I have found that not to be the case and this major technological
breakthrough has permitted me to analyze the effect of blasting and
develop unique product and stability control blasting methods based
upon this scientific discovery.
Livingston Crater Disclosure and Experiments
My early experiments were conducted in crater blasting of many
materials, including ice, snow, clay and a variety of ores. A
crater blast must be distinguished from a bench blast, which is
discussed later. A crater blast is a blast which is detonated
beneath a surface that extends laterally in all directions beyond
the point where the surrounding material would be effected by the
blast. Conversely, a bench blast takes place beneath the surface
that is stepped at a point sufficiently close to the charge that
the vertical face of the step is affected by the explosion.
In analyzing crater blasts, I determined that there is a definite
relation between the energy of the explosive, as represented by the
weight and type of explosive, and the volume of material that is
effected by the blast of such explosive, and that this relationship
is significantly effected by the placement of the charge.
This scientific approach to the art of blasting and blast-mining is
based on the scientific concepts which I have discovered and which
form the basis for a new technology. This technology permits
analysis of mining problems by a systematic engineering evaluation
as opposed to the previous approach employed by industry of relying
on experience and hit-and-miss blasting and has resulted in the
method of this invention.
In order to understand the significance of the new technology
disclosed herein and the associated control blasting and mining
method, it is appropriate that the history of my experiments be
summarized and the analysis that was required to reduce the data
gained from these experiments to scientific principles.
From numerous blasting experiments, I determined that a
strain-energy relation exists, which can be defined by the
equation:
N=E.cuberoot. W (1)
wherein N, termed the "critical depth," is the depth in feet at
which displacement of the surface above an explosive charge of
spherical shape does not exceed a specified limit in response to a
blast with the type and weight of explosive in a given homogeneous
material beneath a horizontal surface of semi-infinite lateral
extension. The limiting displacement is that at which cratering of
the horizontal surface begins and the term "semi-infinite lateral
extension" represents a surface sufficient that vertically arranged
lines about its extremities are not effected by the blast;
wherein E, termed strain-energy factor, is a variable factor
dependent upon the explosive and the material, but for the same
explosive and material is a constant for charge weights in the same
order of magnitude when cube root scaling is assumed; and
wherein W is the weight of the explosive charge in pounds.
This equation may be rewritten as:
d.sub.c =.DELTA. E.cuberoot. W (2)
wherein .DELTA. equals d.sub.c /N and d.sub.c is the depth in feet
from the surface of a material to be blasted to the center of
gravity of the explosive charge. Thus, .DELTA. is a dimensionless
term which expresses the ratio of the depth of any charge as
compared with the critical depth N.
When d.sub.c is such that the maximum volume of material is
completely ruptured, d.sub.c = d.sub.o, which is the optimum depth.
The optimum depth d.sub.o is that depth of charge which creates
maximum pressure rise in the material, and is determined by the
maximum coming of the material, and is that depth at which
fracturing proceeds to completion, but no further.
The distinction between d.sub.o and N can be best illustrated by
reference to FIGS. 1 and 2.
FIG. 1 illustrates a crater blast in material of semi-infinite
extension with the explosive charge placed at critical depth N. As
is illustrated by the concentric lines, the shock wave created by
the explosion radiates outwardly, to form a spherical shock wave,
passing through the material, and the largest or outside circle
indicates the limit of participation of any materials in response
to the explosion. In effect, this charge is placed at such a depth
that the effect of the charge will just reach the surface of the
material in which the explosive is set, and ground rise will not
exceed that at which cratering begins, but ground rise may be
sufficient to cause some cracks to form. N is the depth at which
the charge will just begin to have a cratering effect, but is not
so close to the surface that any significant crater will actually
be formed.
Referring to FIG. 2, the effect of an explosion set at optimum
depth is illustrated. The charge is placed at optimum depth d.sub.o
from the surface of the material, and again the shock front which
extends in spherical relation radially about the charge is
illustrated by concentric circles in progressive stages of time.
When the shock front reaches the surface of the material, the
reflection of the shock at the free face as a factor of time is
illustrated by the series of arcuate lines projecting down from the
surface, and this reflection of the shock produces slabs of what is
commonly called "fly rock." The extent of doming due to massive
ground motion is illustrated by the hemispheric line raised above
the surface. This doming characteristic of massive ground motion is
produced at a later time by the pressure rise in the material as
distinguished from the pressure at the shock front. Referring to
FIG. 3, a representative pressure time chart is illustrated, which
distinguishes the difference between the pressure due to the shock
and later pressure due to molecular vibration. FIG. 4 shows the
slabbing effect of arrival and reflection of the shock front as
distinguished from maximum breakage per pound of explosive caused
by maximum pressure rise within the material. When this ground
rise, or doming, reaches maximum extent without escape of the gas
bubble formed by the explosion, the energy of the explosion has
been used to its ultimate; and in effect the greatest volume of
material for a given charge has been broken. The difference between
the effect of a blast resulting from a charge placed at critical
depth and of a blast having a charge placed at optimum depth is
this: The surrounding material effected by a charge placed at
critical depth is not isolated by the cratering process, although a
portion may be completely fractured. Rather, the material is
weakened in some areas, although not completely broken, whereas the
material effected by the blast of a charge placed at optimum depth
has a significantly greater portion broken to recognizable loading
and unloading fracture limits and forming a substantially
homogeneous mass of a different texture.
The critical depth and optimum depth for any given explosive and a
given material are determined by test crater blasts (i.e., the
blasts are conducted in a material having a semi-infinite
horizontal free face) and a number of test blasts may be required
to determine these values for the given explosive and the
material.
In performing these test blasts, several charges are placed at
varying depths in bore holes and then each charge is detonated to
determine the values for d.sub.o and N. The effect of an explosive
and the manner in which the test blasts are conducted is
illustrated in FIGS. 5 through 8.
FIG. 5 illustrates the effect of a blast which is detonated above
the surface of the medium and shows the apparent crater formed by
such a blast. FIG. 6 illustrates the result of a blast from a
charge positioned at surface level and shows both the apparent
crater and the true crater formed by such a blast. FIGS. 7 and 8
show further examples of the effect of different depths for the
same charge weight on both the apparent crater and the true
crater.
FIG. 9 illustrates what is meant by the terms "apparent crater" and
"true crater," as well as other terms that are employed in
connection with this description of the reaction of surrounding
material in response to a blast. The apparent crater is the
elevation of the broken rock within material below the original
ground surface in response to the blast, while the true crater is
the limit at which the broken rock is completely isolated from the
surrounding material.
The limit of extreme rupture is a surface running through the ends
of the fractures farthest from the center of the explosion. This
extreme rupture limit contains material effected by the blast
insufficiently to separate it from its surrounding material, but
may contain a variety of types of cracks.
By understanding the effect of a blast of explosives on the
surrounding material, it is readily apparent that it is preferable
to utilize the energy of the explosive in most instances to create
the maximum true crater so that the material effected by the blast
is crushed to the maximum possible extent and thereby its
rehandling facilitated. If the larger portion of the energy is
directed to producing an extreme rupture limit of considerable
thickness, then use of the energy is not maximized.
As previously stated, the value of d.sub.o for a given material and
a given explosive is determined by test blast. I have determined
that the optimum depth is that depth which produces maximum
pressure rise within the material, and that maximum pressure rise
occurs when maximum doming without venting is achieved. In
understanding this concept, it is important to realize that an
explosion creates, in addition to the shock effect, an explosion
cavity including a gas bubble which swells and expands through the
material. When the depth of the charge is such that the surface of
the surrounding material rises to a hemispherical shape, but no
further (i.e., doming), maximum pressure rise within the material
is attained and the gas bubble just breaks through to the surface
of the surrounding material. Thus, as can be seen in FIG. 2, the
response to a blast set at optimum depth is a swelling of the
surrounding material with the surface of the material expanding to
hemispherical shape and the gas bubble created by the blast just
breaking through the surface. Of course, maximum pressure rise is
substantially reached when the depth of the charge is such that the
gas bubble does not break through the surface of the material, but
would require only de miminum additional pressure to cause such
breakage. When the gas bubble is lost to the atmosphere,
recoverable elastic potential energy stored during the rise to
maximum pressure is released so as to produce unloading type
fractures at the walls of the crater.
During an explosion, the disturbance consists of a "front" or shock
front. An abrupt pressure rise occurs at this front, followed by a
series of waves that cause a pressure rise within the material
behind the shock front. It has been previously felt that when an
explosion is detonated sufficiently close to ground surface such
that the shock waves are not dissipated sufficiently during their
travel through the medium as to be dampened out before reaching the
surface, the movement of the surface is brought about by reflection
of the shock at the free face. While this is true to some extent, I
have found that the maximum energy utilization of an explosive
creates what I term "massive ground motion," which effects a much
greater volume of material than that due to the arrival and
reflection of the shock and is produced by the pressure rise in the
material behind the shock front and by reaching critical frequency
of vibration of the particles themselves.
When the charge is detonated, the shock front radiates from the
point of detonation and an explosion cavity containing a gas bubble
is created within the medium. If the point of detonation is
sufficiently close to the ground surface, ground rise is produced,
which results in some fracture of the surface material. Upon an
even closer relation of the charge to the surface, the gas bubble
breaks through the surface, vents to the atmosphere and dissipates
the energy of the explosive still remaining within the gas bubble,
and creates a secondary explosion. The pressures defining the gas
bubble radiate outwardly from the point of breakthrough so that the
surface material that has risen with the explosion is fractured and
projected in all areas to evidence the common occurrence of fly
rock.
Two methods of determining d.sub.o are available.
One utilizes a technique in which V/W (volume of material broken
per pound of explosive) within limits of complete rupture is
plotted against the depth ratio, .DELTA., to determine the specific
value of .DELTA..sub.o, where V/W is maximum. The other combines a
technique including direct observation of ground rise to determine
which charge depth produces maximum doming and pinpointing to
determine critical depth. High speed motion pictures may be used in
this method to assist in determining the velocity and duration of
ground motion and when the gas bubble escapes from the
hemispherical doming produced by a charge at optimum depth. A depth
d.sub.o, the energy utilization number A is unity and the maximum
proportion of the total energy of the explosion is expended in (1)
fractures formed during the rise to peak pressure and (2) fractures
formed during implosion or elastic rebound by release of potential
energy stored during loading.
In either event, it is usually necessary to conduct a number of
test blasts at various depths in order to determine the exact depth
that is optimum. If a number of charges of constant weight are
placed at a range of depths approximately at optimum depth, it is
possible to observe the doming in response to the explosion and
plot a curve based on depth of charge versus extent of doming and
use this curve to determine the optimum depth.
Since cube root scaling is employed in evaluating the test blast,
the optimum depth for any weight of explosive may be calculated
from the parameters gathered during the test blast which are
performed with a comparatively small explosive charge. Cube root
scaling is a system of scaling in which is assumed the following
preposition:
N.sub.1 /N.sub.2 = d.sub.1 /d.sub.2 = .cuberoot. W.sub.1
/.cuberoot. W.sub.2 ( 3)
wherein d.sub.1 is a given distance from an explosive charge of the
weight W.sub.1, and d.sub.2 is the distance from a second charge at
which the same damage effect is felt from the second geometrically
similar charge of weight W.sub.2. Material and explosive type must
be the same in both instances. However, this scaling, which has
proved in practice to be based on a valid assumption, permits the
art of blasting to be more properly analyzed as a science.
The equations discussed above and associated parameters are
summarized in the breakage process equation which is definitive of
the new technology that I have discovered and is written:
V/W = E.sup.3 ABC (4)
this breakage process equation can be used in understanding and
establishing desired blasting effects for any given material with a
given type explosive. The terms V, W, and E have been previously
defined and the additional terms A, B, and C are defined as
follows:
A is the ratio of volume of material broken in the true crater for
a charge of given weight set at any depth as compared with the same
charge set at optimum depth A = V/V.sub.o.
B is the "material behavior index" defined by the formula B =
V.sub.o /N.sup.3 and is constant for a given explosive, a given
material and a given weight of charge. It is a dimensionless number
that may be thought of as a ratio of volumes at two different
energy levels, one when the energy is utilized to a maximum and the
other when the energy is used less efficiently.
C is the stress distribution number which is a dimensionless number
that depends upon the energy distribution within the material and
the duration of the action. It is included as a ratio of the volume
excavated for any given condition to the volume excavated at the
reference condition.
C stands for charge shape, one of the principal factors affecting
stress distribution for the geometry of crater blasts, and through
extensive test in crater blasting it has been found that the most
efficient charge shape is spherical, so that the spherical charge
shape has been used as a definitional reference. Thus, in crater
blasting, C defines more specifically the effect of a charge of
given weight and shape at a given depth as compared to the effect
of a charge of spherical shape under similar conditions.
From these equations, it can be seen that the maximum utilization
of energy of an explosive charge can be determined for any given
type of explosive in a given material. Since cube root scaling is
employed, a number of test blasts can be fired to determine the
parameters of the equations, and these parameters can then be used
for any other charge weights for determining the associated optimum
depth.
For example, if the optimum depth for a 10-pound spherical charge
in a given material is found by test blast to be 5 feet, and the
associated critical depth is found to be 10 feet, then E can be
calculated from the formula N = E.cuberoot.W.
e = n/.cuberoot.w
e = 10/.cuberoot.w
e = 10/2.15
e = 4.65
and .DELTA..sub.o = d.sub.o /N = 5/10 = 0.50
Assuming now that it is desired to employ a charge of 1,000 pounds
of a spherical shape in the same type material, the optimum depth
may be determined:
d.sub.o = .DELTA..sub.o E.cuberoot.W
d.sub.o = 0.5 .times. 4.65 .times. .cuberoot.1,000
d.sub.o = 0.5 .times. 4.65 .times. 10
d.sub.o = 23.2 feet
As a further example, assume the tests have given the same
parameters and that it is desired to find the charge weight, that
is optimum for a blast at 100 feet.
d.sub.o = .DELTA..sub.o E.cuberoot.W
100 = 0.5 .times. 4.65 .times. .cuberoot.w
.cuberoot.w = 100/2.325 = 43
w = 79,500 pounds
Understanding these relations of explosives and the reaction of
surrounding materials to the blast of a given charge, it is now
possible to evaluate and scientifically analyze blasting
applications. Control of the results produced by a blast
pre-designed and placed to achieve such controlled results is now
possible.
The background material discussed to this point is directed
primarily to crater blasting, but the true significance of the
invention is the adaptation of the principles recognized in
conjunction with crater blasting to the art of bench blasting, and
the method that I have devised for controlling the trajectory and
granulometry of the material surrounding the explosive charge in
conjunction with bench blasting.
Accordingly, it is a primary object of this invention to provide a
method for controlling the trajectory, granulometry and stability
of the material surrounding an explosive charge in performing bench
blasting.
It is another object of this invention to maximumize energy
utilization of an explosive charge in bench blasting.
Still a further object of this invention is to provide a method for
determining the optimum placement and weight of an explosive charge
in bench blasting to control the trajectory, granulometry and
stability of the material surrounding the charge.
A still further object of this invention is to provide a mining
process that employs trajectory control, product control and
stability control blasting to significantly reduce the cost per
yard of ore mined.
These and other objects of the invention will become apparent to
those of ordinary skill in the related art when consideration is
given to the accompanying detailed description of the preferred
embodiment of the method of the invention and steps in practising
the invention along with the examples provided and when such
consideration is taken in conjunction with the accompanying
drawings.
DESCRIPTION OF THE DRAWINGS
FIG. 1 illustrates the effect of an explosive charge placed at
critical depth;
FIG. 2 illustrates the effect of an explosive placed at optimum
depth;
FIG. 3 illustrates the time-pressure variation for a charge in the
range d.sub.o following an explosion;
FIG. 4 illustrates the slabbing effect resulting from arrival and
reflection of the shock front;
FIGS. 5-8 illustrate the various crater shapes resulting from
charges placed at various depths;
FIG. 9 illustrates the material variation in response to an
explosion and the terminology employed;
FIGS. 10 and 10a are representative of A vs. .DELTA. charts;
FIGS. 11, 12 and 13 illustrate the various fracture patterns
resulting in rock in response to an explosion;
FIG. 14 is a chart showing representative curves of scaled slant
distances vs. scaled charge depths for the various fracture
patterns;
FIG. 15 is a scaled chart of curves based on scaled slant distances
vs. scaled charge depths for the fracture patterns and having a
superimposed predominate particle size chart;
FIG. 16 is a chart showing N-scaled crater depths plotted against
.DELTA. ;
FIG. 17 is a blast pattern design for autogenous blasting;
FIG. 18 illustrates an autogenous blast design for a large cell
mining unit;
FIGS. 19a-19c illustrate an exemplary N-scaled .DELTA..sub.b
product control blast design;
FIGS. 20a-20c illustrate an exemplary N-scaled .DELTA..sub.v
product control blast design;
FIG. 21 is an exemplary N-scaled autogenous product control blast
design;
FIG. 22 is a typical preferred particle size or percent passing
chart;
FIG. 23 is a plan schematic of the autogenous blast design of FIG.
21;
FIGS. 24-27 illustrate various stages of the mining method in
progress; and
FIGS. 28-31 illustrate various aspects of the land reclamation and
hydropower development phase of the invention.
DETAILED DESCRIPTION
Cratering Parameters and the Breakage Process Equation
Trajectory control and product control blasts are "bench type"
blasts in which two or more free faces occur, as distinguished from
crater blasting in which a single free face occurs. A typical bench
blast usually contains several blast holes detonated in a
predetermined sequence and delay interval, whereas a typical crater
blast consists of a single blast hole.
Although I have previously published articles on the cratering
theory and have written the Breakage Process Equation in its
general form, I have not previously disclosed the fundamental
relations between crater and bench blasting nor have I described
the design technology of trajectory control and product control
blasting.
To provide a background for the new technology of bench blasting, I
will first relate some of the discoveries that made it possible for
me to progress from cratering technology previously described to
the technology of trajectory control, product control and stability
control blasting for bench geometry. Of the various stages of
development of my work, perhaps the most significant was the
realization that "massive ground motion" and motion of the ground
due to reflection of the shock wave were two different things. The
relation between massive ground motion and motion of the ground due
to the reflection of the shock wave became more apparent as my
research was directed towards bench blasting.
Massive ground motion occurs later than the arrival of the shock
wave at the free face and is closely associated with the pressure
rise in the material. I have discovered that the pressure rise
within the material due to blasting depends upon the charge depth
and is maximum at d.sub.o.
The depth at which cratering begins does not necessarily coincide
with the depth at which fracture or cracking begins. The critical
depth is related to the charge weight by the previously stated
equation:
N = E.cuberoot.W
The strain-energy factor E for the explosive, the material, and the
scale of the experiment is constant and is computed using observed
critical depth N and the known charge weight W. Similarly, using
the equation .DELTA..sub.o = d.sub.o /N and the charge depth
d.sub.o at which the presure rise is maximum, I can find the
numerical value of .DELTA..sub.o at which the pressure rise in the
material is maximum. Table 1 summarizes values of E and
.DELTA..sub.o that I have determined for various explosives and
materials.
I have discovered that the volume of material broken per pound of
explosive (V)/W is maximum for a given explosive and a given
material at .DELTA..sub.o in crater blasting. This relation makes
it possible to determine the charge depth at which the pressure
rise in the material is maximum without placing pressure guages in
the material.
Now it so happens that venting of the gas bubble begins at
precisely the instant and charge depth at which the volume of
material broken per pound of explosive in crater blasting is
maximum. Accordingly, it is possible at optimum depth to determine
the time to massive ground motion using a wire wrapped around a
charge which breaks the circuit at the instant of detonation and a
wire along the surface by which the circuit is broken at the
instant at which ground motion begins.
It is necessary to know the time to massive ground motion when
blasting a number of sheet-like benches in rapid succession. In
multiple row blasts, for example, the ground projected by
detonation of the first row of holes must be displaced from the
position in front of the next bench so as to expose the new free
face. Accordingly, detonation of the second row must be delayed at
least the length of time to massive ground motion. This is
important because the motion of the ground from the second row of
holes to be detonated must not be dampened by the pressure rise
from the first row of holes, so that breakage limits due to any
given hole or combination of holes is predictable, or so that the
trajectory of a given row is not adversely influenced by the
preceding row of holes.
To adjust for the geometry of the bench, I first get values of E
and .DELTA..sub.o for the crater condition. Test blast in cratering
conditions also gives me the numerical value of B, the materials
behavior index, where
B = V.sub.o /N.sup.3
by measuring the volume of the true crater from an explosion at
d.sub.o and performing the previously described test to determine
N, where V.sub.o is the volume of the crater in cubic feet at
.DELTA..sub.o, and N.sup.3 = E.sup.3 W
By definition, for the geometry of a prototype crater blast using a
charge of spehrical shape, values of A and C of the Breakage
Process Equation both are unity hence, the Breakage Process
Equation V/W = E.sup.3 ABC reduces to V.sub.o /W = E.sup.3 B for a
blast at charge depth d.sub.o and for a charge of spherical shape.
Under these conditions, A = 1 and C also = 1.
The numerical value of the energy utilization number A of the
equation V.sub. o /W = E.sup.3 ABC may be determined for cratering
type blasts by measuring the actual crater volume V for a spherical
charge of a given type and weight of explosive at any given charge
depth d.sub.c, as compared to the volume measured at d.sub.o.
I have determined the variation in A, the energy utilization
number, for many types of explosives when blasting in a variety of
materials and have discovered that there are two basic A vs .DELTA.
curve shapes: one for brittle materials and one for plastic acting
materials.
I have also discovered that the points of inflection (points at
which a change in slope of the curves occur) corresponds with the
charge depths at which a change in the partitioning of energy from
the explosive to the material occur. Further, upon experimenting
with various types of explosives in a given material and in various
types of material with a given explosive, I have discovered that
the explosive, the geometry, and the material are not separable,
independent variables. Instead, they are dependent variables. This
means that the fundamental variables in blasting are energy, mass,
and time. It also indicates that there is a relationship between
materials behavior in blasting and the theory or relativity.
The effect of the change from crater geometry to bench geometry is
to decrease the mass to which the energy of the explosion is
partitioned, and also to decrease the time to reach massive ground
motion. Applied to bench blasting, this means that the distance
from the charge to the bench free face must be increased as
compared to the distance to the horizontal free face, because of
the gravity forces on the material above the charge that are not on
the material to the side of the charge, if the pressure rise within
the material is to be maximum and the velocity of massive ground
motion is to be the same as for crater geometry. The geometry of
the bench blast differs from that of a crater blast in that two or
more free faces are exposed -- one at the top of the bench and one
at the face of the bench, possible one or more in the hole spacing
direction depending upon the blasting sequence. The geometry of the
bench, and hence the mass of materials to which the energy of the
explosion is partitioned, differs from the cratering condition.
I have assigned a value of unity to A, the energy utilization
number at optimum depth for crater blasting and at optimum burden
for bench blasting. The term "burden" (d.sub.b) indicates the
perpendicular distance from the bench free face to the charge,
optimum burden occurs when the charge is at a burden distance from
the free face to effect maximum energy utilization. The numerical
value of B remains constant, and is determined for cratering
conditions. However, values of B can be used in the Breakage
Process Equation as applied to bench blasting so long as the crater
and bench blasts are conducted with the same explosives, in the
same material, and are of the same magnitude.
The only parameter that changes in the Breakage Process Equation
for bench blasting at optimum burden and depth compared with crater
blasting at optimum depth is the stress distribution number C. I
have found that the numerical value of C for a blast of given
charge shape and type of explosive varies with the bench geometry
as compared to crater geometry as well as with the shape of the
explosive charge.
If a single charge bench blast is detonated at a vertical depth
d.sub.o using a spherical charge of given weight and type of
explosive, the burden distance must be increased to a value greater
than the value of the distance d.sub.o to compensate for the
decrease in confinement and the fact that the material in the
vertical direction is subject to a greater gravity pull than is the
material in the horizontal direction. The burden at which the
breakaway velocity is the same in the free face direction as for a
cratering blast at .DELTA..sub.o is called "optimum burden,"
d.sub.bo. The numerical value of .DELTA..sub.bo (optimum burden
ratio = d.sub.bo /N ) in the burden direction always is greater
than .DELTA..sub.o and the difference .DELTA..sub.bo -
.DELTA..sub.o is called "the delta increment." The numerical value
of increment (.DELTA.inc) is determined by test blasts, as is also
the velocity of ground motion at optimum burden. The .DELTA.
increment should not be confused with the .DELTA. difference
(.DELTA.diff) which is discussed later.
At optimum burden in bench blasting, the .DELTA. increment
(.DELTA..sub.bo -.DELTA..sub.o) depends upon the explosive and the
material, but usually is in the range of 0.02 to 0.10.
When blasting with a long column of explosive, either in crater
blasting or in bench blasting, the energy distribution is effected.
It may be impractical in conventional mining practice to use
charges of spherical shape; if so, tests are conducted using the
hole diameter and charge shape to be used in practice. The value of
C thus obtained always is less than the true value for prototype
conditions (i.e., spherical shaped charges).
When blasting with multiple charges detonated simultaneously, a
reinforcement occurs which tends to increase C. Prior to the bench
blasting, hole spacing cratering blasts are detonated
simultaneously at optimum charge depth to determine the optimum
spacing between adjacent charges d.sub.so which gives a depth ratio
.DELTA..sub.so = d.sub.s /N at which V/W is maximum. The value of
d.sub.so is determined by crater blasting tests with side-by-side
charges detonated simultaneously. The charge spacing for a given
weight of explosive which produces the maximum volume of broken
material is d.sub.so and is converted to the spacing ratio by
dividing by N (i.e., .DELTA..sub.so = d.sub.so /N ). The spacing
ratio is thus a scaling factor which permits one to determine the
optimum spacing for any given weight charges by the formula
d.sub.so = .DELTA..sub.so E.cuberoot.W
When changing to bench geometry blasting, the d.sub.so can be
scaled using the parameters from the cratering test. The spacing
d.sub.s may be adjusted from the value d.sub.so at the option of
blast designer in order to control the breakaway velocity of the
bench free face. Moving the adjacent charges closer together
increases breakaway velocity and separating the charges decreases
the velocity.
I have found that the optimum hole spacing depends upon the
explosive and the material and is greater than the optimum depth.
Usually, .DELTA..sub.so ranges from 1.05 to 1.2 .DELTA..sub.o for
charges of spherical shape. Long, cylindrical charges are less
effective and require closer spacing. The exact value of
.DELTA..sub.so is determined for the explosive, the material, the
drill hole geometry, and the bench geometry to be used in
practice.
The optimum stress distribution number C.sub.o, for bench blasting,
identifies the conditions at which maximum pressure rise occurs in
the material and at which massive ground motion begins. The
condition occurs at optimum burden, optimum charge depth and
optimum hole spacing for bench blasting. It takes into account the
shape of the explosive column and the reinforcement of adjacent
charges. The prototype charge is spherical, and adjacent charges
are detonated simultaneously. A change from spherical to
cylindrical shape results in a decrease in volume broken per pound
of explosives. Similarly, a departure from simultaneous detonation
to delay detonation also results in a decrease in volume broken per
pound of explosives at optimum pressure rise.
The numerical value of C.sub.o is evaluated using the Breakage
Process Equation. At optimum burden, optimum depth and optimum hole
spacing: A = 1.0, and B= (V.sub.o /W ) crater .sup.. (1/E.sup.3).
At optimum bench geometry and optimum pressure rise within the
material, C = C.sub.o and the Breakage Process Equation reduces
from
(V/W) bench = E.sup.3 ABC.sub.o
to
(V/W) bench = (V.sub.o /W ) crater .sup.. C.sub.o
from which
C.sub.o = [(V/W) bench]/[(V.sub.o /W) crater]
where
(V/W) bench is the volume of material broken per pound of explosive
in bench blasting at optimum pressure rise, and
(V.sub.o /W) crater is the volume of material broken per pound of
explosive in cratering blasting at optimum pressure rise.
The optimum stress distribution number C.sub.o may be thought of as
a conversion factor from crater blasting to bench blasting. More
specifically, it may be thought of as relating the ratio of volumes
broken in bench blasting and in cratering blasting for a charge of
given weight, type and shape at that instant at which massive
ground motion begins. At the reference condition where massive
ground motion begins and the pressure rise in the material is
maximum, stored elastic potential energy is released from the
material, and the momentum of massive ground motion attributable to
the explosion also is maximum. The momentum of massive ground
motion equals the sum of the masses of all particles of the
material broken times the sum of the breakaway velocities of all
particles participating in ground motion.
Referring now to the Breakage Process Equation and recognizing that
C.sub.o identifies a specific condition (whereas C does not) it
follows that a bench blast might "liberate" either more or less
rock than identified by the condition of energy partitioning at
C.sub.o. The liberated rock may or may not be within limits of
fracturing due to the blasting failure process.
It follows from the preceding discussion that the volume of
material capable of being broken per pound of explosive is greater
for bench blasting than for crater blasting, due to the decrease in
confinement. It is important to recognize that volumes are compared
at that condition of energy partitioning at which the pressure rise
in the material is maximum. It is also important to recognize that
the optimum volume in bench blasting occurs at optimum burden,
optimum depth and optimum hole spacing where the velocity of ground
motion equals that for crater blasting at optimum depth.
The determination of .DELTA..sub.so and C.sub.o involves both
crater and bench blast tests. The volume of material broken per
pound of explosive at optimum pressure rise in the material for the
geometry of bench blasting is computed from the bench geometry at
optimum charge depth, optimum burden and optimum hole spacing, and
the value of C.sub.o the optimum stress distribution number is
computed from the equation
C.sub.o = [(V/W) bench]/[(V.sub.o /W) crater]
or from the equation
C.sub.o = V/N.sup.3 B
It is important to know the optimum stress distribution number for
bench blasting, because it provides a means of referring cratering
performance to bench performance. The term "performance" should be
thought of in relation to molecular behavior, energy partitioning
and energy distribution rather than in relation to the volume of
material excavated.
Consider for a moment the form of the Breakage Process Equation
(V/W) bench = E.sup.3 ABC
The left side of the equation may be thought of as a measure of the
ratio of the mass of material broken to the energy of the
explosion. Similarly, the right side of the equation also must
represent a mass to energy relationship. The equation which defines
B may be written in the form:
B = (V.sub.o /W) crater .sup.. (1/E.sup.3)
substituting the value of B into the Breakage Process Equation
there results
(V/W) bench = (V.sub.o /W) crater .sup.. A .sup.. C
which states that for a given explosive, a given material and a
charge of given shape, the volume of material broken per pound of
explosive in bench blasting equals the volume broken per pound of
explosive in crater blasting times the AC product. The energy
utilization number is defined by the equation
A = (V/V.sub.o) crater
V = crater volume at charge depth d.sub.c
V.sub.o = crater volume at optimum depth d.sub.o
Numerical values of A depend upon energy partitioning and vary with
the depth ratio. It is difficult, if not impossible, to describe
energy partitioning in absolute units. The method I have devised is
based upon relative units with the standard reference at
.DELTA..sub.o. It is assumed that the volume of material broken by
a given explosive charge of spherical shape is a measure of energy
partitioning to rock breakage. FIGS. 8 and 9 illustrate the
variation of A with .DELTA. as determined for various types of
explosives and materials. It may be observed that the depth ratio
at which V/W and pressure rise at that instant of massive ground
motion are maximum depends both upon the explosive and the
material. Observed and experimentally determined cratering
parameters for various explosives and materials are summarized in
Table 2.
Breakage Limits and Type of Failure
Three types of failure occur in blasting; they are (1) shock
failure, (2) shear failure and (3) viscous damping failure. Shock
failure is characteristic of brittle-acting substance, such as
granite and anorthosite. Shear failure is characteristic of more
ductile materials, such as altered rocks, soft iron ore and ice.
Viscous damping failure is characteristic of composite substances,
such as snow, which contain air in the voids. Both the solid
portion (snow crystals) and the gas portion (air) influence the
failure process.
Understanding the mechanics of the blasting failure process is
important to the understanding of product control blasting. The
mechanics of the blasting failure process differ for each of the
three known types of failure. Hence, the slant distance from the
center of the explosive charge to the end of a given type of
fracture formed by the explosion also differs. FIG. 9 shows these
slant distances which are important in determining the controlling
factors of product control blasting. The slant distance from the
center of gravity of the explosive charge to the rim of the true
crater is C.sub.r, and to the limits of extreme rupture is
C.sub.e.
Fractures may be classified into three main groups -- (1) R
(radial) cracks which are vertical planes that radiate outwardly
from the bore hole, (2) O (onion) cracks which form concentric
shells resembling the form of the wave front, and (3) I (inward)
cracks which appear on the surface as concentric circles
surrounding the bore hole and dipping inwardly toward it and are
generally conical. The R, I, and O cracks are mutually
perpendicular and are oriented parallel to the three principal
strain axes -- one oriented in the radial longitudinal direction,
the other two in the spherical surface of the wave front.
Schematics of these different types of fractures are shown in FIGS.
11, 12, and 13. FIG. 11 illustrates the radial cracks which are
shown radiating outwardly at 1 from the vertical bore hole as lines
2 extending from the center of gravity of the charge. Reproduction
of the effect or appearance of the radial cracks is difficult, and
it must be understood that FIG. 11 is schematic only and the cracks
are shown by the lines of the drawing, the surrounding material
defining the cracks not being illustrated for sake of clarity.
FIG. 12 illustrates the O cracks which are designated 3 and are in
the form of stacked sphere surface segments perpendicular to the
axis of the bore hole or to a vertical line passing through the
center of gravity of the charge. This illustration is exploded
relative to the other crack illustrations for the sake of
clarity.
FIG. 13 illustrates the I cracks. These cracks outline a number of
frustroconical sections in stacked relation with their axes
perpendicular to each underlying layer of onion cracks.
Of course, it should be understood that these cracks are all formed
in response to a single blast and are superimposed on the
surrounding material. The types of cracks have been separated only
to show them more clearly. Too, the onion cracks are illustrated
relatively larger than the R and I cracks for clarity.
Crater formation is intimately associated with the cracking system
defined by the principal strain axes, but may include shearing type
fractures in which the failure surfaces are conical making an acute
angle with the radial direction, or a shock type fracture in
brittle-acting materials in which slabs are formed in successive
layers parallel to the free face. Cratering of ductile material is
accompanied by uplift and doming of the surface.
The R, I and O cracks, the shear fractures in ductile materials,
and the slabs in brittle materials are formed during the rise to
peak pressure and may be classed as loading type fractures.
Unloading type fractures also may develop parallel to the walls of
the excavation during the decline from peak pressure, and a
secondary type of ground motion may occur within the crater due to
migration of the high velocity gases from the explosive. The
secondary type of motion is referred to here as the "vortex action"
and control of vortex action may be used to supplement control of
the primary massive ground motion in product control blasting where
abundant fine size material is required.
The extent to which the fracture pattern develops and doming occurs
in a given material depends upon the energy of the explosion,
physical properties of the material and the depth of the charge.
FIG. 14 identifies the relation between scaled slant distance to
various breakage limits and scaled charge depth for sandstone
conglomerate, using a high velocity explosive. The shape of the
curves describes the variation of breakage limits with .DELTA. and
is indicative of partitioning of energy to the material, or of
venting of the gas bubble, which marks the beginning of loss of
energy to the atmosphere. The Breakage Limit Curves are determined
by energy partitioning and are functions of the A vs .DELTA. curve.
In the example, doming develops to a lesser distance than radial
cracks, but the reverse may be true in other materials. In other
than brittle-acting materials, where shock type slabbing
predominates, the R cracks extend outwardly to the greatest
distance, the I cracks next, and the crater rim terminates to a
lesser distance than the I cracks.
FIG. 10 is typical of the variation of A with .DELTA. in frozen
iron ore when blasting with ANFO. Limits of fracturing and
deformation are divided into ranges and subdivided into regions
with limits defined by various key values .DELTA..sub.d,
.DELTA..sub.ss, .DELTA..sub.f, .DELTA..sub.o, and .DELTA..sub.a. At
charge depth .DELTA..sub.d = 1.4 doming begins. The region between
.DELTA. = 1.4 and 1.0 is referred to as the uplift region.
Cratering for the shearing type failure characteristic of frozen
iron ore begins at .DELTA. =1. The region between .DELTA. = 1.0 and
0.87 is the shear region or "isolation" region of the doming range.
The isolation region is one in which the granulometry of the broken
material is coarse, but becomes less coarse as .DELTA..sub.f is
approached. The crater radius is maximum at .DELTA..sub.f and for
frozen ore when blasting with ANFO maximum crater radius occurs at
.DELTA..sub.f = 0.87, which marks the completion of the primary or
"loading-type" crater fracturing process.
The crater increases in volume from .DELTA..sub.f to .DELTA..sub.o
which is referred to here as the primary region of the
fragmentation range due (1) to the change in crater shape during
development of loading type fractures and (2) to unloading type
fractures which begin at .DELTA..sub.f and increase the volume of
the crater to the abrupt change in slope near the top of the A vs
.DELTA. curve (about .DELTA.=0.85 in this case). At .DELTA..sub.o
=0.83 the gas from the explosion (gas bubble) vents to the
atmosphere and energy remaining within the gas bubble is
partitioned largely to the atmosphere. Venting marks the beginning
of the secondary region of the fragmentation range, which for the
blasts of ANFO in frozen iron ore ends at .DELTA..sub.a = 0.30
where the volume of the apparent crater is maximum. A blast within
the secondary region of the fragmentation range becomes
increasingly inefficient from the standpoint of the volume of
broken material per pound of explosive. It also becomes
increasingly hazardous due to uncontrolled throw of the flyrock and
because of the airblast pressure in the vicinity of the explosion.
Although inefficient from the standpoint of energy utilization,
three additional types of fragmentation predominate with the
secondary region of the fragmentation range (1) fracturing due to
destruction of bond strength in a zone close to the charge; (2)
fracturing due to the vortex ground movement of the rock as it
implodes inwardly towards the explosion cavity and upwardly into
the rising column of gas and flyrock; and (3) fracturing due to
impact of rock falling upon the ground. The airblast range begins
at .DELTA..sub.a and continues into negative values of .DELTA.. At
.DELTA. = 0 the center of gravity of the charge is at the ground
surface and d.sub.c = 0.
Cratering type test blasts are conducted to evaluate parameters E
and B of the Livingston's equations, to measure breakage limits and
the variation of A vs .DELTA., and to identify limits and
sub-limits of the doming range, fragmentation range and airblast
range. As the work proceeds, the velocity of flyrock travel, the
velocity of massive ground motion and the time and duration of
ground motion are measured.
Single crater blasts are supplemented by two hole and three hole
crater blasts to determine the optimum hole spacing and to evaluate
the "reinforcement" effect of simultaneous detonation upon crater
volume and upon the velocity of massive ground motion at
.DELTA..sub.o. Using the crater data, bench blasts are designed at
optimum charge depth and optimum hole spacing to determine the
optimum burden and to evaluate C.sub.o. The velocity, time, and
duration of massive ground motion are measured with a high speed
camera.
Preliminary blast pattern designs for trajectory control and
product control blasting are made using Livingston's equations and
cratering and bench tests employing design principles later
discussed in detail. The preliminary designs finally are tested at
full scale utilizing such field instrumentation as may be
desirable, then are modified if necessary, to improve the economics
of drilling.
Elements of Trajectory Control Blasting
The method of trajectory control blasting makes it possible to
predetermine the breakaway velocity of the bench blast and to throw
the material in a preferred direction. Control of trajectory may be
accomplished so as to throw the major portion of the material
horizontally, vertically, or in an inclined direction. If thrown
horizontally, the trajectory is called a .DELTA..sub.b trajectory.
If thrown vertically, the trajectory is called a .DELTA..sub.v
trajectory. If thrown inclined upward at 45.degree., the trajectory
is called a maximum range trajectory. A maximum range trajectory
becomes a maximum economic range trajectory if the material breaks
away at an upward inclination of 45.degree. with a velocity
approaching that obtainable for a charge at optimum burden, optimum
depth, and optimum hole spacing.
Control of trajectory depends upon controlling massive ground
motion rather than upon the control of particles liberated by
arrival and reflection of the shock wave. Massive ground motion of
a bench blast occurs with maximum momentum at optimum burden,
optimum depth, and optimum hole spacing, but the trajectory cannot
be controlled under these conditions. To control trajectory
requires use of a "buffer wall." A buffer wall is a mass of
material located so as to inhibit ground motion in a given
direction. The buffer wall delays the time and reduces the duration
of ground motion in a direction other than that of the desired
trajectory and reduces the velocity but does not necessarily
eliminate the motion. The buffer wall may exist in space at the
instant of blasting, or it may be created during the action of the
blast. The thickness of the buffer wall is identified by the
.DELTA. difference between the free face towards which massive
ground motion is inhibited and that in the direction of the
controlling burden toward which it is promoted. If .DELTA..sub.b is
the .DELTA. value in the near horizontal burden direction, and
.DELTA..sub.v is the .DELTA. value in the vertical direction,
ground motion will be promoted vertically if .DELTA..sub.b exceeds
.DELTA..sub.v. The .DELTA. difference controlling the direction of
massive ground motion will be .DELTA..sub.v - .DELTA..sub.b. As the
.DELTA. difference increases, the buffer wall remains intact for a
greater period of time and less massive ground motion occurs in the
unwanted direction.
Not only is it necessary to control the .DELTA. difference, but it
is necessary also to control the .DELTA. value in the direction of
desired ground motion. For example, for a horizontal trajectory
.DELTA..sub.v must exceed .DELTA..sub.b and the value of
.DELTA..sub.b determines the breakaway velocity. If .DELTA..sub.b
is greater than .DELTA..sub.bo the breakaway velocity will be less
than that at maximum economic range assuming the .DELTA. value in
the spacing direction = .DELTA..sub.so. Conversely, if
.DELTA..sub.b is less than .DELTA..sub.bo the breakaway velocity
will be greater than that at .DELTA..sub.bo, but the momentum of
massive ground motion will be less than maximum due to the loss of
energy that escapes from the break-through of the gas bubble prior
to the time when maximum pressure rise occurs. The decrease in
momentum occurs because the mass in motion is reduced more than the
velocity is increased.
I have observed from numerous blasting trials that a cratering
charge at .DELTA..sub.o, or a bench blast at optimum burden
.DELTA..sub.bo domes up the surface as a hemisphere equal to the
doming radius at .DELTA..sub.o. Accordingly, for the condition of
optimum burden, optimum charge depth, and optimum hole spacing, the
breakaway velocity of massive ground motion equals
v = .sqroot.2gh
v = breakaway velocity, ft/second
g = the acceleration of gravity, and
h = the doming radius at .DELTA..sub.o
Using a breakage limit drawing such as constructed from the crater
test, we may determine the doming radius and compute the breakaway
velocity. The breakage limit chart is prepared by "scaling" the
slant distances from the center of gravity of the charge to the
extreme limits of doming for a number of charges and plotting these
limits against the scaled depth of the charge. For example, using
FIG. 14 for test blasts in sandstone conglomerate with a high
velocity, high energy explosive, the cube root scaled doming radius
at .DELTA..sub.o equals 3.6 and the doming radius for a 1,000 pound
charge would be 3.6 .sup.. .cuberoot.1,000 = 36 feet. The velocity
of ground motion would be v = .sqroot.2gh = .sqroot.64.4 .sup.. 36
= 48 feet per second. Thus, if the direction of the ground motion
is known by design of the free face and use of the buffer wall, the
placement of the material can be determined using the known
velocity of the ground movement.
In addition to the relation between doming radius and breakaway
velocity at .DELTA..sub.o, I have observed from numerous blasting
trials that in the region of optimum burden, optimum depth, and
optimum hole spacing the breakaway velocity changes along the face
as the slant distance increases. The breakaway velocity also
decreases as the charge depth increases. The relation describing
these changes is in the form
v.sub.2 /v.sub.1 = r.sub.1.sup.2 /r.sub.2.sup.2
where
v.sub.1 is breakaway velocity at .DELTA..sub.o
v.sub.2 is breakaway velocity at desired burden
r.sub.1 is scaled charge depth at .DELTA..sub.o
r.sub.2 is scaled charge depth at new desired burden
If, as a continuation of the previous example, a breakaway velocity
of 60 ft/second is required in the direction of the controlling
burden, it would be necessary when blasting on a bench of given
height in sandstone conglomerate to reduce the scaled burden from
2.05 at .DELTA..sub.o to 1.835 computed as follows:
60/48 = (2.05).sup.2 /r.sub.2.sup.2
giving
r.sub.2 = 1.835
If the reduction in burden were accomplished at constant
.DELTA..sub.v, the thickness of the buffer wall would be increased,
and more of the material would be thrown in the burden direction,
but the momentum of massive ground motion would decrease. To
achieve an increase in breakaway velocity and yet minimize the loss
of momentum the designer might elect to hold the thickness of the
buffer wall constant. By doing so, however, he would not recover
all of the lost momentum of massive ground motion, because the
numerical value of A in the breakage process equation for the new
conditions would be less than 1.
In a highly simplified form, the method of trajectory control
blasting can be summarized as follows.
First, it is necessary to determine the values of N, E, d.sub.o,
.DELTA..sub.o, d.sub.bo, .DELTA..sub.bo, d.sub.so, and
.DELTA..sub.so (if more than one charge is to be blasted along the
line of the free face). These determinations must be made for the
particular explosive and material to be blasted.
These values are originally determined by test blast. Values for N,
E, d.sub.o, and d.sub.so are found from crater blast. The values
for d.sub.bo are determined from test bench blast. These test bench
and crater blasts are made with small charges of the explosive in
the material in which full scale blasting is to conducted.
Values for .DELTA..sub.o, .DELTA..sub.bo, and .DELTA..sub.so are
calculated by dividing d.sub.o, d.sub.bo, and d.sub.so respectively
by N.
The second step is to prepare the free face of the bench so that
the angle of trajectory will be dictated by the .DELTA. difference
and the angle with the horizontal made by the line connecting the
charge center of gravity perpendicularly with the closest free
face. Preparation of the free face is usually done by the design of
a preceding blast. The velocity of breakaway of the bench free face
is next calculated from the formula v = .sqroot.2gh with the h
value taken from a chart showing the scaled doming radius plotted
against the scaled charge depth, which chart is made from
measurements of phenomena occurring during the crater test
blast.
The .DELTA. difference depends also upon requirements of product
control and is discussed more in detail in connection with product
control blasting.
FIG. 15 shows the slant distances and the controlling factors for
the .DELTA. difference. Some of these factors are:
1. .DELTA. difference fixes the back break limit and predetermines
.alpha. for the next row of holes.
2. .DELTA. difference determines the breakaway velocity in the
unwanted direction. As .DELTA. difference increases velocity in
unwanted direction and duration of action decreases.
3. .DELTA. difference determines particle size in the buffer wall
zone. As the .DELTA. difference increases, the particle size in the
buffer wall zone increases and the volume of the particles of
larger size also increases.
4. The .DELTA. difference must exceed .DELTA. increment; otherwise
ground motion is at random.
The .DELTA. difference can be determined geometrically to gain the
desired back break of the material. If the angle of trajectory
.alpha. is to give maximum range, it will be maintained at
45.degree. so that the angle of the free face with the horizontal
must be 45.degree.. Thus, if it is desired to continue with a
maximum range trajectory, the back break will require a c.sub.r
slant line of 45.degree. with the horizontal. In order to achieve
this, the value of d.sub.v will be calculated from the geometric
relation as d.sub.v = c.sub.r cos .alpha.. The value of d.sub.b
will be dictated by the desired breakaway velocity and .DELTA.
difference numerically established as .DELTA..sub.b -
.DELTA..sub.v.
My earlier experiments had led me to believe that by placing the
explosive charge at a depth of .DELTA..sub.o, and a distance from
the mid-point of the free face, less than .DELTA..sub.bo,
trajectory in the direction of the free face could be insured.
However, I have now discovered that this analysis is incorrect or
at least incorrect to the point that maximum utilization of the
energy of the explosive charge is not available in such an
arrangement. Rather, I have now found that more efficient control
of the trajectory of the material is gained by placing the
explosive charge a distance equal to .DELTA..sub.bo from the free
face and a distance greater than .DELTA..sub.o from the horizontal
face of the material. This, in effect, supplies a buffer area of
material above the area in which maximum pressure rise within the
material is attained, whereby maximum pressure rise within the
material is gained to the point coinciding with the vertical free
face with respect to the explosion and its spacing from the free
face. In this regard, I attain maximum utilization of the energy of
the explosive charge and get maximum pressure rise and ground rise
which generally means optimum velocity of the mid-point of the free
face when the explosive charge is placed a distance of d.sub.bo
from the mid-point of the free face. While I do not get maximum
utilization of the energy with respect to the horizontal surface or
upper surface of the surrounding material due to the buffer zone,
the buffer zone does operate to limit the movement of the material
in response to the explosive charge and thus inhibits the ground
rise in the direction of the upper surface of the material while
permitting the maximum ground rise in the direction of the free
face. Accordingly, by this placement of the explosive charge, the
maximum trajectory of the surrounding material may be attained in
the free face direction. The exact angle of trajectory, as
previously described, is controlled by the angle of the free face
and its relation angularly within the position of the charge.
In the trajectory control blasting, a number of charges along a
free face may be desired in order to blast a considerable length of
overburden along a free face from its in situ position into its
final resting position. An additional factor must be considered in
this situation, since there is considerable reinforcement of one
charge by the detonation of an adjacent charge. The adjacent charge
is not of sufficient strength so that the placement between charges
can be equal to twice the value of .DELTA..sub.o. The distance
between adjacent charges is given a value .DELTA..sub.s, and
normally to attain maximum utilization of the energy in a
trajectory control blast, the spacing between the adjacent charges
.DELTA..sub.s is equal to somewhere in the range of 1.2 -
1.4.DELTA..sub.o.
In this spacing of charges, and the reinforcement effected by
simultaneous blasting of the adjacent charges, an increase in
utilization of the explosive energy is gained.
Elements of Product Control Blasting
1 -- Introduction
Product Control Blasting includes .DELTA..sub.b Product Control
Blasting, .DELTA..sub.v Product Control Blasting and Autogenous
Blasting, and is a blasting technology in which the size of the
broken material and the range in granulometry in bench blasting is
controlled.
The Breakage Process Control Mining Method utilizes .DELTA..sub.b
and .DELTA..sub.v Product Control Blasting when excavating
overburden to prepare material for various zones of the terraces,
embankments, and pillar dams. The method also utilizes autogenous
blasting of the orebody to prepare a suitable feed to the
processing plants.
The technology involves the use of one or more or various
combinations of:
a. Breakage limit control
b. Ground motion control
c. Bond strength control, and/or
d. Vortex action control
The granulometry of the broken material may be controlled so as to
produce:
1. predominately coarse material
2. predominately fine material, or
3. a graded product of specified size and gradation.
Although the prior art recognizes a relationship between the
quantity of explosives used and the size of the broken rock,
Product Control Blasting provides a means of pre-determining
particle size and of controlling the granulometry of the broken
material and represents a substantial technological breakthrough.
The new technology overcomes a series of erroneous concepts
generally held under the present state-of-the-art.
For example, it is generally held that if predominately fine
material is to be produced, small diameter blast holes must be
used. Similarly, it also is held that the granulometry of the
broken material is determined primarily by physical properties of
the material or by geologic conditions such as the spacing between
joints, rather than the design of the blast. As a result of my
research, I have discovered that material as weak as snow can be
blasted into large blocks, that a new fracture pattern can be
superimposed upon the existing pattern governed by the geology, and
that rock can be broken by blasting to a predetermined size and
range in granulometry.
Under the present state-of-the-art of mining, it is assumed that
the broken rock from a mine must pass successively through a coarse
crushing circuit, an intermediate circuit, thence to fine grinding.
However, with autogenous blasting as disclosed herein, a major
portion of the ore can be reduced in a single step from run-of-mine
size to finished product without the use of crushing machines. In
autogenous blasting, as in an autogenous mill, the ore acts as the
grinding media. In mining operations where conventional crushing
and grinding circuits are employed, autogenous blasting can be used
to reduce the overall cost of drilling, blasting, haulage, loading,
crushing and grinding. This is possible because the low cost energy
of explosives properly harnessed is capable of improving the output
of loading, haulage and crushing equipment and of reducing labor
costs and capital investment in mining equipment.
As further example of the need for Autogenous Blasting, first
consider the growing tar sand industry. Due to the tremendous rate
at which raw ore must be moved from mine to processing plant and
the large volume of material that must be handled in a short time,
it becomes economically impractical to screen, stockpile, or crush
the material ahead of the processing plant. In my view, Autogenous
Blasting, which has the capability of producing a product of
pre-determined size and range in granulometry, either substantially
reduces or completely eliminates the need for screening,
stockpiling or crushing, and therefore provides a solution to a
difficult problem.
In the application of Autogenous Blasting to oil shale mining, the
raw material either may be fed to a TOSCO type retort, which
requires a fine feed, or to a gravity type retort which accepts a
graded product somewhat coarser in granulometry, but at the present
stage of development, experiences difficulties of channeling and
clinkering in the retort. The problems of feed preparation for oil
shale mining--including the harder material, the more stringent
feed requirements for the retorts than for tar sand separation
drums, and the even greater feed rates to the processing
plants--are perhaps even more acute than for tar sand mining.
.DELTA..sub.b Product Control and .DELTA..sub.v Product Control
Blasting supplement Autogenous Blasting and also are of economic
importance because they are capable of more economically producing
a graded product of coarse size such as rip-rap and break water
stone than by conventional methods. Under conventional methods, the
yield of stone meeting specifications usually is low, hence an
excessive volume of material must be excavated to obtain a given
quantity of usable material. Many instances can be cited under the
present state of-the-art of the difficulty of economically
producing a suitable graded product for the rock shell of a rolled
rock-fill dam, or for filter and slope protection material for
earth-filled dams. Product Control Blasting substantially minimizes
this difficulty.
The economic advantages of Product Control Blasting and Autogenous
Blasting and the probable effect of the technology of the invention
upon the mining and construction industries should become readily
apparent from the description in the following pages.
2 -- The Problem and the Breakthrough
Rock failure in blasting occurs in a definite, predictable sequence
which is determined by the energy-mass relations. A first step
towards control of the blasting failure process was my discovery
that the energy/mass relations of blasting could be controlled so
as in a series of blasts to duplicate each of the steps of a
sequence for a single blast that might occur within the space of a
few milliseconds.
The next step was to apply what I had learned about crater blasting
to the new geometry of bench blasting, so as to control bench
blasting results. Having learned how to control the trajectory of a
bench blast, I concluded that it also should be possible to control
the blasting failure process so as to predetermine the size of the
broken material and the range in granulometry. As my work
proceeded, I came to the conclusion that by combining several blast
holes each of predetermined size and relation to each other that
almost any type of gradation required in mining and construction
could be designed for and produced. The steps which allow me to
achieve this desired result and effect product control constitute
this phase of the invention.
At this stage of the work, the search first was directed toward
identifying design parameters and describing in numbers the
variation in particle size when bench blasting with a single hole.
Later, as the technology developed, the bench was directed toward
determining how to convert data from a single hole bench blast to a
multiple hole blast.
FIG. 15 summarizes results of a key stage of the work, which is
this instance is illustrated for a given rock type, (Pottsville
Sandstone) and for a given explosive (Hercomite 2). FIG. 15 shows
the variation in breakage limits with charge depth for crater
blasts graphically. Referring to FIG. 15, the scaled charge depth
for a blast is plotted against scaled slant distance determined in
the field for each of several breakage limits (radial) cracks,
second ring I-cracks, first ring I-cracks and crater limits). Key
.DELTA. values (.DELTA..sub.a, .DELTA..sub.o, .DELTA..sub.f,
.DELTA..sub.ss, and .DELTA..sub.d, which can be converted to
numerical values from the formula d.sub.c /.sqroot.W = E .sup..
.DELTA. ) are plotted for reference, because previously I had
learned that changes in energy partitioning occur at these key
.DELTA. values. I found from the breakage limit-particle size
studies that the mechanics of the blasting failure process and
energy partitioning could be broken down into various
characteristic ranges and further sub-divided into regions
representative of the granulometry. I found that:
1. The uplift region of the doming range was characterized by
"pre-condition" of the rock into large blocks not completely
isolated by the blast.
2. That the shear region of the doming range, in which
shearing-type fractures are formed, is characterized by particles
of coarse size -- or large "pie-shaped" blocks that are isolated
from the material.
3. The primary region of the fragmentation range identifies
completion of the primary fracture pattern and development of a
secondary pattern due to the release of stored potential energy
from the material -- it is characterized by a progressive decrease
in particle size with decrease in .DELTA..
4. The secondary region of the fragmentation range is a region of
premature breakway that follows venting; it is characterized by two
extremes -- fine material near the blast hole and coarse material
near the free face.
5. The air blast region of the fragmentation range is one in which
airblast following venting is excessive, and much of the rock is
scattered about the countryside.
When attempting to predict for a multiple hole bench blast the
granulometry that results from blasting a series of holes each
producing its individual gradation, it would be helpful to have a
screen analysis for each type of blast hole. Unfortunately, it is
impractical at the blast pattern design stage either to set up a
screening plant or to recover all of the flyrock from a blast.
Inasmuch as it is impractical to go directly to a blasting-screen
analysis correlation, I concluded that it was more important to
identify the interdependence of granulometry and .DELTA. than to
attain screen analysis curves at various values of .DELTA..
My solution to the prediction of the granulometry that results from
a multiple hole bench blast was to:
1. Extend the crater tests to identify the "predonderant particle
size" at known .DELTA.-values, or within various characteristic
regions of energy partitioning.
2. Relate preponderant particle size to the bond strength
destruction zone, the intermediate zone, and the buffer -- limits
of which are controllable by design.
3. Develop a fundamentally sound technique wherein preponderant
particle and zone volume for a single crater blast could be
converted to preponderant particle size and zone volume for a
single hole bench blast.
4. Develop an analytical technique for the prediction of gradation
curves for a multiple hole bench blast using single hole bench
blast preponderant particle size and bond strength destruction data
for each of three identifiable energy/mass zones of the
borehole.
5. Provide a means of correlating the crater and bench granulometry
design data with screen analysis curves obtainable during the
production stage of blasting.
One of the problems is to estimate the volume and size of the "fine
material" associated with a blast; it cannot be measured in the
crater and cannot be sampled. I concluded, that the fine material
is produced by bond strength damage, which is a function of .DELTA.
and depends upon parameters of the breakage process equation. The
volume of bond strength destruction is less difficult to calculate
than it is to measure -- so I combined size and volume measurements
of the preponderant particle size of material from the intermediate
and buffer wall zones which I could measure, with volume
calculations of bond strength damaged material.
Fine material is here defined as that smaller than some minimum
size practical to sample by hand using a grid network at the
surface of the muckpile (e.g. sand). The bond strength destruction
zone is a zone of rock surrounding a blast hole and within which
the material is crushed and reduced to fine material as above
defined. The predominate particle size is defined here as the
numerical average size of all particles measured using a grid
network for sampling of the surface of the muckpile, and consists
of particles larger than the minimum size as defined for fine
materials
Now to review the steps leading to control of bond strength damage
in blasting and to the method of calculating the volume of fine
material.
The depth of the true crater, when blasting with a charge of
spherical shape, extends beyond the center of gravity of the charge
some radial distance which is referred to here as the "gas bubble
radius." Strictly speaking, the term gas bubble radius equals the
radius of the explosion cavity, but I found that limits of bond
strength damage in "brittle-acting" materials and cavity limits in
"plastic-acting" materials are inter-dependent. Referring now to
FIG. 16, it may be observed that the gas bubble radius is greater
for shallow than for deep charges. The gas bubble radius decreases
to a minimum at .DELTA..sub.o with increase in depth of charge and
remains nearly constant at depths greater than .DELTA..sub.o. The
maximum pressure in the medium due to detonation of an explosive in
a bore hole decreases with the radial distance from the bore hole.
The decrease in pressure with distance for a given material and a
given explosive is called the pressure decay law for that
material.
Although the radius of the gas bubble and the limit of bond
strength destruction depend upon the pressure decay law, they
depend also upon the charge depth and the duration of the action.
Rather than to compute the pressure at a given scaled distance or
to correlate the pressure in the material with the radial distance
from the bore hole, but which methods are available using pressure
gauges, I found it simpler to use a prospector's pick and measure
the radial distance from the center of the bore hole along the wall
of the excavation at which the rock became strong enough so that it
no longer could be readily excavated with a pick. In crater blast
tests, the radial distance is measured at the elevation of the
center of gravity of the explosive charge to the wall of the true
crater. Using the radial distance as measured for a cylindrical
charge such as used in practice, and assuming bond strength
deterioration to occur within a cylindrical zone surrounding a long
column of explosive, the volume of bond strength destruction is
computed. The volume of material too fine to include within the
grid network measurements for predominate particle size thus is
estimated with reasonable accuracy.
The "preponderant particle size" is measured by constructing a grid
work at the surface of the muckpile and sampling particles larger
than say three inches by measfuring the length, width, and depth of
each particle touching the grid line. The volume of the
preponderant particle then is computed as a numerical average
assuming particles to be rectangular, or of the same shape. In
addition, the sampled rock also can be sized by hand and each
fraction weighed so that preponderant particle size and screen
analysis gradation can be correlated, say at the controlling burden
intended to be used. Volumes of preponderant size material and the
fine material are added mathematically to obtain the total volume
of the broken material.
Referring now the lower portion of FIG. 15, which plots
preponderant particle size vs d.sub.c /.cuberoot. W we may observe
that the preponderant particle size gradually increases in the
uplift region of the doming range with decrease in charge depth to
a maximum size near .DELTA..sub.ss (See .DELTA. value where
cratering begins). Further decrease in charge depth causes the
particle size to decrease to a minimum at .DELTA..sub.o where the
pressure rise in the material is maximum. If the controlling burden
is decreased to less than .DELTA..sub.o, the blast falls either in
the secondary region of the fragmentation range or in the air blast
range and the preponderant particle size once again increases as
indicated in the figure. I consider these changes to be consistent
with my observation on energy partitioning, and attribute this
increase in particle size at depths less than .DELTA..sub.o to
"premature breakaway," because massive ground motion occures before
the optimum magnitude and duration of pressure is reached in the
rock. The observed particle size-energy partitioning relations are
consistent with fundamentals of trajectory control blasting in that
maximum pressure rise within the medium occurs at .DELTA..sub.o and
the interval between zero time and the time of massive ground
motion is maximum at .DELTA..sub.o. The term premature breakaway
thus indicates to me that the energy of the explosion is
substantially greater than that at the energy concentration at
which V/W is maximum.
Referring now to FIG. 16, it may be observed that at least a part
of the excess energy is utilized to enlarge the cavity as the
confinement is decreased. The remainder is lost to venting and
causes a decrease in the volume of material broken per pound of
explosive and an observations in flyrock travel distance and
airblast pressure. Energy lost to the atmosphere is not available
to the material -- hence, the particle size at charge depths less
than .DELTA..sub.o increases.
At first I interpreted the relations of FIG. 15, wherein the
particle size increases with decrease in charge depth, at less than
.DELTA..sub.o to indicate that it was theoretically impossible to
reduce the proponderant particle size to less than that at
.DELTA..sub.o. This is the condition that limits conventional bench
blasting in open-pit mining and which results in primary crushers
being required to handle the coarse material that is not size by
blasting.
It was at this stage of my research that I conceived the idea of
Autogenous Blasting. I had discovered that a vortex action
immediately preceded venting, and I concluded that it was practical
to control the direction of the vortex just as I had previously
controlled the direction of the trajectory, and in doing so would
modify energy partitioning. The vortex action is an effect of the
escape of energy from adjacent the charge blast. In effect, the
material surrounding the blast forms a venturi through which the
broken rock is drawn by the reduced pressure of the escaping
material.
As my research continued, I found that by directing the vortex into
the medium rather than into the air, I could suppress both the
airblast and the flyrock hazard and utilize the energy explosion to
do useful work, provided that the depth of the charge was
sufficient and the height to diameter ratio of the cylindrical
charge did not exceed about eight to one. While promoting the
vortex action, I found also that I could increase the volume of
bond-strength-affected material relative to the total volume broken
just as would be expected by preventing energy loss to the
atmosphere.
These observations indicated to me that particle size could be
reduced and blasting could safely be conducted in the air blast
range -- something previously impossible to do. I consider these
observations to be of extreme practical importance. Further
research provided the proof I needed that air blast and flyrock
could be controlled in the air blast range and material could be
crushed finer by blasting than normally is done by conventional
practice. I choose the term autogenous blasting to describe the new
technology, because the attrition accompanying the vortex action
scrubs bond-strength-damaged material to grain size. Without the
controlled vortex action and attrition, moderately bond
strength-damaged material otherwise would retain its cohesion. A
principle of autogenous blasting design therefore is to produce a
well-developed vortex by avoiding "confused" ground motion.
Confused ground motion is ground motion in all directions or in a
random direction rather than in a direction directly toward the
free face. Avoidance of confused ground motion involves:
1. control of the depth of the breakaway point, and
2. control of the vortex pattern of several blast holes so as to
avoid interference one with the other.
3 -- Basic Technology
The technology of Product Control Blasting consists of:
1. Relating particle size and breakage limits in cratering to the
depth ratio .DELTA..
2. Adjusting breakage limits and particle size of crater blasting
using the .DELTA.-increment adjustment method to compensate for the
geometry of the bench in bench blasting.
3. Blasting in the .DELTA..sub.v direction to break horizontal
slabs of rock and achieve coarse fragmentation, or blasting in the
.DELTA..sub.b direction to break geological rock formations of
basically vertical slab configuration. The product control blast
may on the other hand be of the autogenous type wherein a vortex
action is produced and a heretofore unachievably fine granulometry
of material is produced. Autogenous blasting also enables one to
control flyrock and minimize air blast.
4. Preparing the bench blast pattern in accordance with
requirements of ground motion, breakage limit, bond strength, and
vortex action control.
5. Combining particle sizes from each of several zones of a blast
hole to find the volumes of each size fraction and combining them
for several holes so as to represent the percentage passing a
theoretical predominant particle size.
6. Determining the "percent passing" each of the predominate
particle size and add or subtract various code blasts to conform
the blasting pattern in accordance with gradation requirements.
4 -- Types of Product Control Blasting
The geometry of product control blasting may be classified into
three main groups as follows:
1. .DELTA..sub.b product control bench blasting in which ground
motion is directed towards the face of the bench.
2. .DELTA..sub.v product control bench blasting in which motion is
directed towards the top of the bench.
3. Autogenous blasting in which a controlled vortex action is
directed horizontally or in an inclined direction toward a third
free face (at a subtantial distance behind the face of the bench)
that is created in time as a result of the blasting delay
sequence.
4. Autogenous modification of .DELTA..sub.b product control
blasting in which the bottom of the bench is undercut by blasting
in the air blast range so as to minimize breakage limits at the
bottom of the bench, thus liberating relatively undisturbed rock at
the top of the bench in particles of a size previously determined
from the geological conditions.
5 -- Product Control Blasting and the ABC Product
Crater test blasts can be fired so as to permit one to measure the
limit of bond strength destruction and the breakage limits and to
define the variation in A with .DELTA.. During the work,
predominate particle size can be measured and, if desired, a screen
analysis made of a sample fraction of the broken material. As a
result of this work, the variation in granulometry with .DELTA. can
be determined and the effect of a variation in explosive type upon
the numerical value of B, the materials behavior index, and upon
granulometry at any desired value of .DELTA. can be measured.
The breakage process equation for a blast with a given explosive
and a given material at a given scale of blasting can be written in
the form:
V/W = E.sup.3 B .sup.. AC
By inspection of the breakage process equation as above written
(since the left hand side must have the same dimensions as the
right hand side) the AC product may be thought of as a mass to
energy ratio just as is V/W.
In this form, the equation suggests to me that the AC product can
be used as an index of particle size, because E and B are constant
and the quantity E.sup.3 B is a constant. A and C are variables. A
varies depending upon the controlling burden, and C varies
depending upon the bench geometry and inter-relations among
.DELTA..sub.b, .DELTA..sub.v, and .DELTA..sub.s. The variation in C
for bench blasting can be related to a reference condition at
C.sub.o where A is unity, using the equation:
C = (V/V.sub.ob) .sup.. C.sub.o
I tested the concept in the field and found that by manipulating
the AC product when blasting with a given explosive in a given
material, I could control the granulometry of the blast in any size
fraction. Later, when expanding my research to autogenous blasting,
I found that due to variation in the rate of energy release and
bore hole pressure achievable with various types of explosives, it
was sometimes advantageous to use two or more types of explosives
in a single blast, or to combine two or more types of explosives in
a single blast hole. Inasmuch as the materials behavior index B is
sensitive both to explosive type and to the material (see Table 2),
I adopted the ABC product as being representative of the general
case and called the ABC product the "Product Control Index."
My experiments have taught me that a high numerical value of the
ABC product identifies a coarse range in granulometry and a coarse
predominate particle size; whereas a low numerical value of the ABC
product identifies a fine range in granulometry and a fine particle
size. Although a given product control index number represents a
range in particle sizes, which in crater and trial bench blasting
can be correlated with a predominate particle size, the product
control index number does not correlate directly with a particle of
given size. Instead, the product control index number provides a
means of increasing or decreasing particle size in bench blasting
regardless of whether or not the predominate particle size has been
measured.
The Vortex Action, Autogenous Blasting and Granulometry
1 -- Measurements
At a certain stage in blasting, the explosion cavity "implodes"
inwardly and is destroyed. This marks the beginning of vortex
motion within the muckpile. Rock is drawn laterally toward the
focal point of the implosion; thence is directed as a jet stream
toward the closest free face. The interior portion of the jet moves
faster than the exterior portion and more of the gas is
concentrated near the center than at the boundaries of the stream.
Particles within the jet stream are predominantly those nearest the
walls of the explosion cavity and are affected to the greatest
degree by bond strength damage. Bond strength damage is the result
of high frequency vibration of the walls of the explosion cavity
due to impact and kinetic energy of the products of combustion.
Thus, fundamental variables controlling bond strength damage are
the energy of the explosion, physical properties of the material
acted upon, and the duration of the action. Although the physical
properties of the material are not within control of the blaster,
the energy of the explosion, the shape of the charge, the depth and
geometry of the vortex, nad the duration of the action are.
To produce a vortex action there must be implosion. Implosion
occurs if the controlling burden is less than .DELTA..sub.o, or
.DELTA..sub.bo and the controlling burden is in the direction of
the free face, but the velocity of the vortex approaches a maximum
at .DELTA..sub.a where the vortex is well developed, and the volume
of the apparent crater is maximum. Thus, the upper limit of
autogenous blasting occurs at .DELTA..sub.o, but the scouring
action is best developed when the controlling burden is set at
.DELTA..sub.a. The lower limit of autogenous blasting is determined
largely by economics, but usually is somewhere near the mid-point
of the air blast region.
To achieve autogenous blasting, the vortex must be directed both
toward the free face and into a muckpile being formed by the blast.
This is done by controlling massive ground motion using a buffer
wall as previously described, and by predetermining the depth and
inclination of the axis of the vortex. There is a limit, of course,
to the magnitude of the .DELTA. difference in autogenous blasting
and to the depth of the vortex action. As an ideal situation,
.DELTA..sub.v should approach or exceed .DELTA..sub.o, otherwise
vertical flyrock may not be sufficiently inhibited and the vortex
action incompletely developed in the direction of the controlling
burden. Typical vortex action control criteria are given in the
design criteria for autogenous blasting, Table 3.
Due to the high velocity jet action, there is a tendency for the
muckpile to become tightly packed. This packing, if not avoided by
proper design of the blast pattern, makes the removal of muck more
difficult. To overcome the packing action, the angle of
inclination, .alpha., of the jet stream with the horizontal can be
controlled to permit the muckpile to bulk rather than to pack. This
is done by progressively increasing .alpha. to a maximum and then
reducing it to a minimum. The technology of a .alpha. control later
is illustrated in sample calculations.
The layout of an autogenous blast differs from that for other types
of product control blasting and from that for conventional open-pit
blasting. It more closely resembles the blasting pattern for an
underground development heading than for a conventional surface
mine. FIGS. 17 and 21 illustrate typical blasting patterns for
autogenous blasting. The blast is opened up at the "cut" area near
the center of the pattern, by a code 1 blast such as in FIG. 21;
after creating the initial opening or cut, ground motion, rather
than being directed toward the top of the bench, is directed toward
the newly created free face at the cut area, which free face is
backed by a muckpile mass being developed by preceeding holes of
the blast. Autogenous blasting requires that a barrier mass be
maintained at the perimeter of the pattern (see FIG. 18). Barrier
masses are successively destroyed during the blasting sequence; but
in order to do so back break must be minimized when blasting holes
of the carrier zone. The code 5 design of FIG. 21 and Table 3, (see
Appendix) illustrates applicable design criteria (.DELTA..sub.rc is
less than .DELTA..sub.v), and the technology of backbreak control.
Ground motion begins at the free face, whereas the vortex action
begins at the explosion cavity. To obtain the most effective
fracturing of the material, I control the vortex so that it occurs
in a line connecting the center of gravity of the explosive charge
and the breakaway point specified by the design criteria.
Autogenous blast pattern design illustrates the technology that
makes it possible to produce a graded product in accordance with
conventional screen analysis specifications. As a result of crater
and bench trial blasts, the limits of bond strength destruction and
the variation of predominate particle size with .DELTA. are known.
Limits of the intermediate zone and of the buffer wall zone depend
upon the .DELTA. difference and the controlling burden.
Accordingly, the volume and predominate particle size from each of
three zones of a blast can be determined and controlled. If one
wishes to produce five times as much rock of predominate size X
than of predominate size Y, five times as many holes with product
control index numbers in the X range than in the Y range are
included in the design.
2 -- The Conversion from Crater to Bench Geometry
The presence of a second free face affects energy partioning and in
turn particle size. The effect of the second free face in bench
blasting is to:
1. Reduce the duration of the action,
2. Modify the pressure distribution relative to that of a cratering
blast,
3. Modify the fracture pattern, the direction of ground motion, and
the degree of fragmentation, and
4. To increase V/W and C.
By definition, C is sensitive to the bench geometry, duration of
the action, and the pressure distribution in the material. Of the
several parameters of the breakage process equation, the stress
distribution number C is the only one influenced by the second free
face.
Fundamentally, the effect of the second free face is to decrease
the mass of material to which energy might otherwise be partitioned
during the interval between arrival of the shock front at the free
face and the beginning of massive ground motion. The decrease in
mass at constant energy is an energy/mass change and such changes
are identified by the depth ratio, .DELTA.. The problem, as
discussed previously, is to determine the change in depth ratio
that is equivalent to the decrease in mass. In its application to
product control blasting, not only is it necessary to describe the
effect of the second free face, but also to evaluate it relative to
the slant distance to breakage limits and to particle size.
The .DELTA. increment (.DELTA..sub.bo - .DELTA..sub.o) is measured
at the reference condition at which the momentum of massive ground
motion is maximum. The reference condition is identified as
occurring at .DELTA..sub.v = .DELTA..sub.o at .DELTA..sub.s =
.DELTA..sub.so, and .DELTA..sub.b =.DELTA..sub.bo. At the reference
condition for bench blasting A = 1, and B = V.sub. o /N.sup.3
crater, by definition. The equivalent mass to be added in the
burden direction to achieve the same breakaway velocity in the
horizontal direction in bench blasting as previously was obtained
in the vertical direction in crater blasting, thus is determined by
test blasting.
When converting from crater geometry to bench geometry, we are
concerned not only with bench geometry at the C.sub.o reference,
but also with the general case where the controlling burden either
may be greater than or may be less than .DELTA..sub.bo . A large
portion of energy of the explosion is lost to the gas bubble at
controlling burden .DELTA..sub.a. Similarly, much of the energy of
the explosion is partitioned to inelastic behavior and seismic
vibrations at .DELTA..sub.ss. Accordingly, when evaluating the
effect of the second free face, energy partitioning also must be
taken into account.
The method I use is called the .DELTA. increment adjustment. I
compensate for the loss of confinement with a theoretical mass that
varies from the maximum at .DELTA..sub.o to 0 at both .DELTA..sub.a
and .DELTA..sub.ss. At maximum size, the theoretical energy may be
visualized as a rectangular slab of thickness equal to the .DELTA.
increment, of length equal to .DELTA..sub.so, and of height equal
to .DELTA..sub.o. As the controlling burden decreases from
.DELTA..sub.o to .DELTA..sub.a, the thickness of the theoretical
mass decreases from the .DELTA. increment to 0. Similarly, as the
controlling burden increases from .DELTA..sub.o to .DELTA..sub.ss,
the thickness of the theoretical mass also decreases to 0. Thus,
the method increases V/W and C and provides for the greatest
adjustment in breakage limits and particle size for the condition
where the pressure rise in the material is greatest. Energy
partitioning automatically is taken into account, and may be
thought of as due to an increase in the AC product.
FIG. 15 illustrates the adjustment as it applies both to breakage
limits and to predominate particle size for .DELTA..sub.b -- and
for .DELTA..sub.v -- product control blasting. Assume for example
that the .DELTA. increment is determined by field test blast to be
0.07 and E has been determined to be 3.30. The value of 0.07 equals
a scaled charge depth of 0.23 computed as follows:
d.sub.c /.cuberoot.W = .DELTA..sup.. E = 0.07 .sup.. 3.30 =
0.23
Since the chart of FIG. 15 is drawn to scale, the value 0.23 may be
scaled up the line at .DELTA..sub.o to the point X and a straight
line connecting the point X to the .DELTA..sub.a and .DELTA..sub.ss
points at zero on the cratering scaled burden portion of the chart,
the scaled value of .DELTA. increment for any .DELTA. value between
.DELTA..sub.a and .DELTA..sub.ss can be taken from the chart.
In the example of FIG. 15, a bench blast due to the presence of the
second free face, to be equivalent to a crater blast at the
.DELTA..sub.o, must have a burden .DELTA..sub.bo = .DELTA..sub.o +
0.07. The .DELTA. increment adjustment of FIG. 15 is identified by
the dimension line ".DELTA. increment adjustment at .DELTA..sub.o."
Thus, the loss of confinement increases the scaled slant distance
to the limit of radial cracking by shifting to the right along the
radial crack curve from B to B'. Similarly, the effect upon the
scaled slant distance to the crater lip is to shift to the right
along the crater curve from C to C'. The adjustment for the second
free face shifts the blast into the primary region of the
fragmentation range and increases the predominate particle size
from 0.75 cubic feet for the geometry of crater blasting to 1.05
cubic feet for the geometry of bench blasting.
Product Control Blast Pattern Design
1 --The Breakage Process Equation Applied to Product Control
Blasting
A high value of the ABC product indicates a coarse product; whereas
a low value indicates a fine product. The problems are: to control
fragmentation by controlling bond strength damage, fragmentation in
the intermediate zone, and fragmentation in the buffer wall zone;
and to design a multiple hole bench blast so as to obtain an
overall product of the desired size and granulometry from a series
of blast holes. This requires that each of the parameters of the
Breakage Process Equation be known for each type (code) of blast
hole in the pattern.
Consider the equation V/W = E.sup.3 ABC and the available data. V/W
depends upon the bench geometry which in turn depends upon the
criteria set for bench design. Design criteria fix the direction of
ground motion, and of the vortex action following implosion, the
limits of bond strength destruction, the breakage limits, and the
size of particle. We may view the resulting fragmentation as
divided into three zones:
1. The bond strength damage zone near the blast hole;
2. The .DELTA. difference zone which acts as a buffer wall to
control ground motion; and
3. The remaining portion between the bond strength zone and the
.DELTA. difference or buffer wall zone.
In effect, the designer, by fixing the limits of each of these
three zones, exercises control over fragmentation within each zone.
The energy concentration is greatest in the bond strength
destruction zone and least in the buffer wall zone. Because
particle size is a function of energy concentration, the buffer
wall zone is the chief contributor to the coarse fraction.
Similarly, the contribution of the fine fraction from the bond
strength destruction zone and of the intermediate fraction from the
intermediate zone is proportional to the ratio of the volume of a
given zone to the total volume excavated by the blast. The volume
from each of the three zones may be computed from the field test
data or from the scaled drawing. Once the design criteria are
stated, the bench geometry is fixed, and V/W can be calculated.
The statement of the criteria depends on the desired predominant
particle size and the values of the parameters found in the test
blast. The test blast provides the parameters for preparing a chart
similar to FIG. 15. Following the steps required to prepare the
.DELTA..sub.b product control design of FIG. 19, one would first
conduct test blast and prepare the chart of FIG. 15. The desired
predominant particle size for the intermediate zone would be
selected and the chart entered at the appropriate point; in this
case, the desired size is 1.70 cu.ft. and entering the chart at
1.70 gives a scaled burden for crater geometry of 2.61. This value
for d.sub.c /.cuberoot.w of 2.61 gives a .DELTA. value of 0.79
calculated as follows:
d.sub.c /.cuberoot.w = E .sup.. .DELTA.
.DELTA.=d.sub.c /.cuberoot.w / E) = 2.61/3.30 = 0.79
This .DELTA. value would be .DELTA..sub. v if blasting in crater
geometry, but to adjust this value for the fact that it is to set
the controlling burden in the free face direction, a .DELTA.
increment must be used to adjust this controlling burden .DELTA.
value. The scaled .DELTA. increment adjustment (from the previously
prepared scaling chart of FIG. 15) at this .DELTA. value is 0.04 so
that .DELTA..sub.bo would be 0.75.
To get the .DELTA..sub.v value, the predominant particle size for
the buffer zone is selected (see Design criteria chart of Table 4),
in this case 3.25 cu.ft. and the graphs of FIG. 15 entered to get a
d.sub.c /.cuberoot.w value of 3.0. This value of d.sub.c
/.cuberoot.w gives a .DELTA. value of 0.91 which in a crater blast
would produce the desired predominant particle size. To adjust for
the lack of confinement due to the bench geometry, a .DELTA.
increment value must be subtracted from the .DELTA. value of 0.91.
The scaled .DELTA. increment value is the distance y on FIG. 15
when compared to the distance which equals 0.07. Thus the value of
.DELTA. increment is 0.03 and the adjusted value of .DELTA..sub.v
is 0.88.
The statement of the criteria can now be adjusted for other factors
and in the example shown, .DELTA..sub.b has been chosen as 0.73
rather than the optimum .DELTA. value of 0.75 in order to achieve
more velocity in the burden direction. In setting the spacing of
the charges, the .DELTA..sub.s value was determined by the geometry
of the plan and section AA (FIG. 19 b) drawings to be 0.79. This
value was determined by deciding that the distance to the midpoint
toe of the bench should be less than .DELTA..sub.f by an arbitrary
value of 0.05 so as to assure that the fracturing would reach the
midpoint toe and the bench would be broken to grade line since
.DELTA..sub.f is the value at which cratering is maximum. Once
these .DELTA. values are known, a scaled bench geometry can be
drawn from which other values may be determined.
In the general case, the bench geometry is described and drawn to
scale using dimensionless .DELTA.-units (preferably as a function
of N), but W, the weight of the explosive is unknown until the
bench height is finally chosen.
The volume of excavation from the Breakage Process Equation is V =
WE.sup.3 ABC = N.sup.3 ABC, and can be computed from the scaled
drawing. For example, referring to the cross section for
.DELTA..sub.b -Product Control Blasting, FIG. 19, V = 0.9038N.sup.3
and the quantity 0.9038 from the equation V = ABCN.sup.3 must be
the ABC product.
The ABC product thus is fixed by the design criteria -- so the
designer needs to know how to establish the design so as to achieve
the desired type of fragmentation.
From the crater tests A, B, and E are known, from the bench tests
C.sub.o and the .DELTA.-increment are known. C is the only
parameter that depends upon the bench geometry rather than upon the
crater geometry -- it can be greater than C.sub.o or less than
C.sub.o.
The controlling burden of a bench blast is controlled by the scaled
charge depth of FIG. 15. It may be described either in units of
d.sub.c / .cuberoot.W or in units of .DELTA., and the units may be
converted one to the other using the general equation in the
form
d.sub.c /.cuberoot.w = .DELTA. E
The controlling burden determines the breakage limits and the
granulometry. The controlling burden also determines the numerical
value of the AC product. Hence, the numerical value of A is fixed
by the bench design criteria and is equal to the value of A from
the cratering tests. The value of C is unknown; but the value of C
can be computed from the equation
C = V/N.sup.3 AB
and the value of the AC product can be computed from known values
of V/W for bench blasting and V.sub.o /W for crater blasting. C can
also be computed from the equation V = N.sup.3 ABC since V can be
calculated from the scaled bench geometry and A and B are known
from the cratering test.
The designer has as his option the selection of the explosive type.
It sometimes is to his advantage to use more than one type of
explosive in a given hole. The numerical value of B are fixed by
the explosive and the material and is determined from the crater
tests.
As illustrated in FIG. 16, the radial distance to the limit of bond
strength destruction is maximum at .DELTA. = 0, decreases to a
minimum at .DELTA..sub.o and remains nearly constant with further
increase in controlling burden. The designer may increase bond
strength damage by selecting an explosive or combination of
explosives having a high value of B, and by blasting at controlling
burdens less than .DELTA..sub.o. Bond strength damage thus may be
controlled within certain limits, provided that the vortex action
and venting also are controlled. Predominate particle size within
each of the three zones may be evaluated after applying the .DELTA.
-increment correction and is as follows:
1. Particle size within the bond strength destruction zone ranges
from grain size to an arbitrary maximum size defined for grid
sampling to determine the preponderant particle size.
2. Preponderant particle size within the intermediate zone is that
determined by the controlling burden, and is determined using
.DELTA..sub.b or .DELTA..sub.v whichever is least.
3. Predominate size within the buffer wall zone is determined by
the burden to the outer limit of the ground motion control buffer
wall.
Referring now to Table 4 the .DELTA..sub.b Product Control blasting
design summary of FIG. 19, the volume of the buffer wall as
previously calculated by a planimeter represents 23 percent of the
total volume and the predominate particle size (after applying the
energy partitioning adjusted .DELTA.-increment correction to
.DELTA..sub.v using FIG. 15) is 3.25 cu. ft. from the buffer wall
zone. Similarly, the bond strength destruction zone is cylindrical
in shape and of radius 0.10N, the volume of bond strength
destruction is
V.sub.b =.pi.r.sup.2 h= 3.14(0.10N).sup.2 . (0.82N) or V.sub.b
=0.0258N.sup.3
The volume V.sub.m of the intermediate zone is
V = V.sub.b + V.sub.m + V.sub.w
V = 0.0258N.sup.3 +V.sub.m + 0.23V
v.sub.m = 0.77V - 0.0258N.sup.3 = (0.77 .times. 0.9038N.sup.3) -
0.0258N.sup.3 where
V = 0.9038N.sup.3 as determined from calculation or a planimeter
measurement on a scaled chart of the blast design,
so that V.sub.m = 0.6959N.sup.3 - N.sup.3 = 0.670lN.sup.3
Now to find V/W and C for the .DELTA..sub.b Product Control cross
section.
Using the Breakage Process Equation
V/W =E.sup.3 ABC
and given
A =0.95 B = 0.47
e = 3.30 abc = 0.9038 v/w = e.sup.3 abc
v/w = 35.94 .times. 0.9038 = 32.48 cu.ft./pound.
The V/W value is necessary in order to evaluate the total cost of
the job.
The value of C may be computed directly from the Breakage Process
Equation as follows:
C = V/N.sup.3 AB = 0.9038N.sup.3 /N.sup.3 (0.95) (0.47) =
0.9038/0.4465 = 2.02
2 -- the Explosive, the Material, and Bond Strength Control
The effect upon Product Control Blasting of a change in type of
explosive usually is less than the effect due to a change in type
of material. However, a change in explosive type when blasting in a
given material effects each of the parameters of the Breakage
Process Equation to some degree. In general, an increase in energy
of the explosive at constant weight increases V/W, but this is not
always true. If more than one type of explosive is to be used
either in a single blast hole or in various holes of the given
blast, it is essential during the crater and bench test blasts to
determine the effect of explosive type and to obtain curves similar
to those of FIG. 15 for each type or combination of types of
explosive intended to be used.
It may be desirable either to suppress bond strength damage in
.DELTA..sub.v - .DELTA..sub.b -- Product Control Blasting if
predominately coarse product is desired, or to promote it in
Autogenous Blasting if a predominately fine product is desired. A
high yield of coarse product usually is best achieved in the shear
region using low energy, low velocity explosives; whereas for
.DELTA..sub.v - .DELTA..sub.b -- Product Control Blasting as
contrasted to Autogenous blasting a high yield of fine product
usually is best achieved in the primary region using high energy,
high velocity explosives. A change in explosive type usually
requires a change in hole diameter, and may also require a change
in bench height to achieve optimum economic benefit and optimum
yield of size-graded product.
The dashed line to the left side of the predominate particle size
curve of FIG. 15 indicates for Autogenous Blasting a continuing
decrease in particle size in the Secondary and Air Blast Regions of
the Fragmentation Range. The decrease in particle size at
controlling burdens less than .DELTA..sub.o is characteristic of
Autogenous Blasting and fundamental thereto and is made possible
by:
1. The geometry of the blast pattern (See FIG. 17);
2. by suppressing premature breakaway and breaking to a newly
created, third free face bounded by buffer means (see FIG. 18);
and
3. By promoting the vortex action and avoiding confused ground
motion following implosion.
3 -- Blast Pattern Design and Granulometry Correlation
FIGS. 19, 20 and 21 illustrate typical blast pattern designs for
.DELTA..sub.b -- Product Control, .DELTA..sub.v -- Product Control
and Autogenous Blasting, respectively. The figures also serve to
illustrate the calculations and the application of the .DELTA.
increment adjustment for .DELTA..sub.b and .DELTA..sub.v bench
blasting to determine breakage limits and particle sizes. The
.DELTA. increment adjustment is unnecessary for autogenous
blasting. The figures are not intended to indicate that a single
design for .DELTA..sub.b -.DELTA..sub.v -- Product Control Blasting
or for Autogenous Blasting is necessary to the invention. Instead,
it is intended merely to illustrate the application of basic
principles some of which apply also too Trajectory Control
Blasting.
Elements of .DELTA..sub.b - .DELTA..sub.v -- Product Control
Blasting are:
1. Ground motion control;
2. Breakage limits control; and
3. Bond strength control.
Autogenous Blasting includes these three elements plus a fourth --
vortex action control.
Ground motion control is governed by the same basic principles as
previously discussed in detail for Trajectory Control Blasting.
Tables 3, 4 and 5, in connection with FIGS. 19, 20 and 21,
illustrate criteria of design for ground motion control for a
single hole for .DELTA..sub.b and .DELTA..sub.v Product Control
Blasting and for a series of holes for Autogenous Blasting. The
figures and associated tables should not be construed to indicate
that .DELTA..sub.b and .DELTA..sub.v product control blasting are
limited to a single hole. On the contrary, multiple hole blasting
patterns for these two types of blasting can be as complex as that
described in greater detail for Autogenous Blasting. The basic
principles are similar, only the chosen criteria changes.
In the .DELTA..sub.b Product Control example of FIG. 19, ground
motion is towards the face of the bench because .DELTA..sub.b is
less than .DELTA..sub.v. Ground motion is not at random because the
.DELTA. difference of 0.15 is greater than the .DELTA. increment of
0.07.
In the Autogenous Blasting example of FIG. 21, ground motion starts
vertically with the crater blast so as first to develop a third
free face at a substantial distance from the face of the bench. The
code 2 blast directs the broken rock in an inclined upwardly
direction towards the cut area without excessive backbreak. The
code 2 parameter controlling backbreak is .DELTA..sub.cr and is
equal to or less than .DELTA..sub.v. The breakaway point of the
code 2 trajectory is below the midpoint of the code 1 excavation
limit. The code 3 blast directs the vortex nearly horizontally
underneath the code 2 muckpile. The code 4 blast directs the vortex
inclined upwardly at 18.degree. so as not to pack the code 3 muck
too tightly and to permit the muckpile to bulk vertically. The code
5 blast directs the vortex upwardly at 32.degree..
FIG. 15, which is the basis for the calculations of FIG. 19,
applies to a given explosive and a given material; hence, similar
breakage limit diagrams must be prepared from trial blasts for
other materials and other explosives. Breakage limit control not
only determines C, V/W, and granulometry for a given blast, but
also determines the geometry of the resulting excavation.
To illustrate that the ABC product is an index of particle size and
that the granulometry of a blast is determined both by the
intermediate zone and by the buffer wall zone, compare criteria and
particle sizes for .DELTA..sub.b Product Control of FIG. 19 and
Table 4 with those of .DELTA..sub.v Product Control of FIG. 20 and
Table 5. Both the .DELTA..sub.b and the .DELTA..sub.v example
designs produce relatively large size material such as might be
required for the shell of a rolled rock-fill dam, or a graded
product having a high yield of coarse material, or for rip-rap or
slope protection. In general, the top of the bench is a larger free
face than the face of the bench and .DELTA..sub.v Product Control
is better adapted to coarse blasting than .DELTA..sub.b Product
Control.
Breakage limits for the .DELTA..sub.v Product Control example are
determined by the criteria .DELTA..sub.v =.DELTA..sub.f = 0.88, to
give maximum crater scaled slant distance, and by the criteria
.DELTA..sub.b = .DELTA..sub.ss = 1.0. These criteria fix the
.DELTA. difference at 0.12, determine C to be 6.48, .DELTA..sub.rc
to be 1.12, V/W to be 40.4, the predominate particle size from the
buffer wall at the face of the bench to be 3.8 cu. ft., and from
the top of the bench to be 3.15 cu. ft.
Principles of breakage limit control illustrated in the figure for
Autogenous Blasting fix the angle of inclination, .alpha., that the
trajectory of the following hole makes with the horizontal. The
inclination of the trajectory and the inclination of the axis of
the vortex are determined by the same fundamentals even though the
vortex action begins later and commences at the explosions cavity
rather than the free face. Charge shape influences both the
trajectory and the vortex. Short columns of explosives are
preferable to long columns because the confinement at implosion is
more uniform and hence there is less tendency at a given charge
depth towards imperfect development of the vortex.
Breakage limit control (i.e., extent of backbreak) for Autogenous
Blasting obviously is of extreme importance. Without it there would
be neither symetrical development of the vortex, controlled
fragmentation, or control of flyrock or air blast. With it:
1. the predominate particle size is successively reduced from 1.0
cu.ft. for the Code 1 holes to 0.03 cu.ft. for the Code 5
holes;
2. the bench is broken to grade;
3. the muckpile is properly bulked so that loading is made easier;
and
4. the barrier wall at the Code 5 excavation limit is straightened
up so as both to minimize the required barrier wall thickness and
later to permit the wall to be systematically destroyed by blasting
after it has served its purpose.
Referring to the example of autogenous blasting illustrated in FIG.
21, the method for determining an autogenous blast pattern to meet
a stated objective will now be described. The design of FIG. 21 is
prepared for the type of material and explosive for which the chart
of FIG. 15 was prepared.
The design of FIG. 21 was prepared to meet the granulometry curve
shown in FIG. 22. As can be seen from the granulometry curve, a
large percentage of the material is desired to be of a relatively
small mesh size. Accordingly, resort must be had to autogenous
blasting, since, as can be observed from FIG. 15, the fine
gradation cannot be achieved with .DELTA..sub.b or .DELTA..sub.v
product control blasting.
The first step is to perform a crater blast to provide a newly
created bench free face behind barrier walls, into which the vortex
of the code 2 charges can be directed.
The code 1 crater blast is chosen for placement at a .DELTA..sub.v
somewhat less than .DELTA..sub.o, in order to achieve greater bond
strength destruction. Since .DELTA..sub.o is 0.68 from the previous
determination, an arbitrary value of 0.61 is chosen for placement
of the charge in the .DELTA..sub.v direction. From this
.DELTA..sub.v value, FIG. 15 can be used to find the limit of
backbreak. The .DELTA. value will have to be converted to d.sub.c
/.cuberoot.w value, where the chart is entered. When the crater
limit line is reached, the corresponding value of the scaled slant
distance is found, which can be converted from its d.sub.c
/.cuberoot.w value to the .DELTA..sub.cr value. These values are
sufficient to show the geometry of the crater formed by the code 1
blast and the predominate particle size can be found from FIG. 15
in a manner similar to that illustrated in connection with the
.DELTA..sub.b product control blast example. In autogenous
blasting, however, the .DELTA. increment adjustment does not have
to be made, since in effect the blasts are of the crater type.
The code 2 blast design must be directed into the muckpile being
formed by the code 1 blast. In order to achieve a smaller
predominate particle size, the trajectory of the code 2 blast must
be directed perpendicular to the free face formed by the code 1 and
the vortex of the code 2 blast directed into the expanding code 1
muckpile. In order to avoid a .DELTA. difference which would
produce a larger range of granulometry, the .DELTA..sub.b value
should be chosen so as to produce a substantially vertical limit of
backbreak. By directing the trajectory of the code 2 blast
perpendicular to the midpoint of the free face minus 0.12 the axis
of the vortex is oriented to minimize surface venting, and from
this selection of the design criteria (see Table 5), the
.DELTA..sub.b value was found to be 0.41, the value being dictated
by the selection of a .DELTA..sub.v value equal to the bench
height. These design criteria from the chart of FIG. 15 result in
100 percent of the material broken being in the intermediate and
bond strength destruction zones and having a predominate particle
volume of 0.10 cubic feet. The spacing between the code 2 charges
is dictated by the geometry of the blasting pattern (see FIG. 22)
and by the amount of material which is desired in the 0.10 cubic
feet predominate particle size range.
The code 3 design is primarily concerned with trajecting the broken
material in a horizontal direction in order to avoid packing of the
previously broken muckpile. The geometry of the code 2 blast which
can be graphically drawn out in the example gives the new free face
into which the material is blasted. In the code 3 design, the
.DELTA..sub.b value is shown equal to .DELTA..sub.a in order to
produce high bond strength destruction. From the geometry of the
layout, the value of .DELTA..sub.v can now be calculated to be
0.71. With this blast design, it will be seen that a
.DELTA.-difference will exist which will produce a buffer wall
zone, the predominant particle size of which can be determined to
be 0.80 cu. ft. from the graph of FIG. 15. The volume of the
material within the buffer zone can be calculated from the geometry
of the blast. The volume of material broken in the intermediate
zone can similarly be calculated so that the predominant size of
the percentage of material which is found to be in the range of
predominate particle volume is 0.25 cubic feet. The selection of
these design criteria establish the geometry of the code 3 blast
and give the volume of material falling within the particular
predominant particle size ranges. The spacing of the holes is
determined by the number of holes selected which, in turn, is
dictated by the geometry of the blasting pattern and the amount of
the material desired in the predominant volume range previously
calculated. Similarly, calculations are made for the code 4 and
code 5 blast designs and the number of charges placed in each
design will be determined by the percentage passing graph shown in
FIG. 22. If the design as laid out in the autogenous blasting
example did not produce the desired range of granulometry, the
number of blasts of any particular code could be adjusted to adjust
the range of predominant particle size produced. For example, were
it found from the calculations previously described that an
insufficient amount of material would be produced to pass 45
percent of the material at 0.25 in the screen analysis, additional
blast charges of the code 3 type could be set, or on the other
hand, if more than 45 percent of the material were passing at 0.25,
the number of code 3 blasts would be reduced. The volume of the
total blast design for this example is shown in Table 6.
The significance of the autogenous blasting method is that the
design layout for a complete mining or other application could be
made prior to any field blasting other than the test blasting to
attain the chart of FIG. 15. The small amount of time required to
make these calculations is more than offset by the tremendous
investment in machinery and crushing costs that would be required
to produce the same range of granulometry if the blasting were
conducted on a hit-and-miss basis.
Bond strength control is not a simple function of explosive
selection or of hole size. Maximum bond strength damage depends
upon developing a maximum impulse within the material at and beyond
walls of the borehole. The pressure rise does not continue beyond
the time of massive ground motion, but the duration of the action
is increased by increasing the depth at which the vortex
occurs.
Bond strength damage may be seen by direct observation, with the
aid of a microscope and polished or thin sections, or by post
blasting strength tests. From blast tests in sandstone with TNT the
limit of complete bond strength destruction was found to extend
outwardly a radial distance from 0.32N at .DELTA. = 0 to 0.15N at
.DELTA..sub.o. In granite the limit extended from 0.27N at .DELTA.
= 0 to 0.12N at .DELTA..sub.o. At controlling burdens greater than
.DELTA..sub.o the limit appears to remain constant and
approximately equal to that at .DELTA..sub.o.
Table 6 identifies limits of bond strength destruction for the
example Product Control Blasting pattern of FIGS. 19, 20 and 21,
and illustrates a typical calculation to determine the volume of
fine material produced by a blast, which otherwise could not be
compensated for. The table also illustrates that the FIG. 19
.DELTA..sub.b Product Control design produces a higher proportion
of fines than does the .DELTA..sub.v design of FIG. 20, and that an
Autogenous Blasting pattern is capable of reducing a substantial
portion of the total volume of material to grain size. Referring to
Table 6, note that nearly 50 percent of the code 5 material and
roughly 15 percent of the total is smaller than the cut-off size
for surface grid sample preponderant particle size
measurements.
The total volume broken by a blast hole equals the sum of the
volumes of rock from the bond strength destruction zone, the
intermediate zone, and the barrier zone. Table 6 records the
volumes of the bond strength destruction zone for each type (code)
of blast hole of the example Autogenous Blasting design of FIG. 21.
Using Table 6, the design summary Table 3, and the plan layout of
the Autogenous Blast of FIG. 21, we may summarize the particle size
distribution from each blast hole and combine all the blast holes
for the Autogenous Blasting pattern to obtain a gradation curve for
the entire pattern.
The left side of Table 7 summarizes the particle size distribution
for each type hole included in the Autogenous Blasting example
Cross Section 21 of FIG. 23. The volume of bond strength
destruction zone for holes, codes 1 through 5 of Table 7, is
obtained from Table 6. The total volume of excavation per hole,
controlling dimensions, and particle sizes from the intermediate
and buffer wall zones of each hole as given in the cross section
and summary tables of FIG. 21 provide all data necessary for the
volume calculations of the left side of Table 7. "Particle Size
Distribution Per Hole."
Using the particle size distribution per hole and a plan of the
blast holes for the Autogenous pattern, we may compute the total
volume of grain size material in the pattern and construct a
predominate particle size distribution curve for the blast. As may
be observed, the right side of the table summarizes preponderant
particle size volumes in order of increasing particle sizes. The
percent passing column is the ratio of cumulative volumes smaller
than a given size to the total volume expressed as a percentage.
The pattern particle size distribution data are presented
graphically in FIG. 22 in a form usually used for gradation curves,
except that "size" is referred to volume rather than to linear
dimensions.
By inspection of the composite predominate particle size gradation
curve (see FIG. 22) for the exemplary Autogenous Blast we may
conclude:
1. A blast hole may be designed so as to predetermine the
predominate particle size in each of the zones of a blast hole, and
so as within certain limits to control the volume of each zone and
the overall granulometry of the hole.
2. A series of blast holes, each of known granulometry
distribution, may be combined so as to obtain a composite
granulometry distribution for the pattern of holes. Within limits
imposed by the technology of ground motion, breakage limit, and
vortex action control, a product control blasting pattern may be
adjusted so as to produce a range of granulometry in accordance
with gradation specifications.
3. If a test blast for which a composite gradation curve is
available is fired, and the resulting muckpile sampled as for a
conventional screen analysis, the proportion of any size fraction
either can be increased or decreased as desired by expanding the
partial plan of FIG. 21 to an overall plan such as illustrated for
a given unit in FIG. 18.
4. composite patterns may be designed for .DELTA..sub.b Product
Control Blasting and for .DELTA..sub.v Product Control Blasting
utilizing the technology described here and illustrated more in
detail for Autogenous Blasting. Similarly, control blocks at the
boundaries of an Autogeneous Blast can be destroyed using
.DELTA..sub.b Product Blasting to produce a product of
predetermined size and range in granulometry.
In 1968 the cost of coarse crushing for large mining operations in
the United States ranged from 7 to 12 cents per ton. The cost of
fine crushing ranged from 13 to 21 cents per ton, and the cost of
loading and haulage ranged from 13 to 21 cents per ton. The
technology of Product Control Blasting makes possible a substantial
reduction in the overall cost of drilling blasting, mucking,
haulage, and crushing and I consider it to be of extreme importance
to the mining and construction industries.
Elements of Stability Control Blasting
Stability Control Blasting is a phase of the technology which I
have discovered which enables one to control the backbreak of
material in response to a blast and, thus, ensure the stability of
the material beyond the backbreak limit. This phase of the
invention is adaptable to the mining process as well as standard
mining procedures in that it permits one to make a substantially
predesigned cut by use of proper placement of an explosive
charge.
Stability Control Blasting is a method and control blasting of
surface and underground excavation having as its objective the
control of static and dynamic elastic rebound. Static elastic
rebound is due to release of elastic potential energy stored in the
rock and releaseable as a result of a change in rock load or a
reduction in lateral restraint. Dynamic elastic rebound is due to
release of elastic potential energy stored in rock due to dynamic
loading for an explosion, a portion of which is recoverable.
The stability control phase of the invention is very closely
related in some aspects to the Product Control Blasting in that it
requires test blasts and production of breakage limit curves
substantially as illustrated and described in connection with FIG.
15. The test blasts required to produce these curves will not be
described again but considering that these preliminary phases of
this aspect of the invention have been completed, the factors and
controlling steps of Stability Control Blasting will be
substantially as follows.
Once the chart of the type illustrated in FIG. 15 is prepared from
the test crater blasts, the basic criterion for Stability Control
Blasting is available in deployment of this criterion to the
specific blast design may proceed. The objects of the Stability
Control Blasting is to break away the surrounding material to a
predefined limit without effecting the stability of the material
beyond that limit. In doing this, what is termed a stability
control zone is selected which is in effect an area adjacent the
material that is desired to leave uneffected in which some
fracturing of the material could be accomplished by the normal
blasting procedure but this normal blasting is controlled to not
produce any fracturing beyond the stability control zone. In
accomplishing this, the control criterion is to assure that the
limit of extreme rupture from the particular charge, weight, and
placement deployed will not extend beyond the stability control
zone. These factors must be realized when normal production
blasting is performed since otherwise, a production blast would
produce fracturing into the area of material that it is desired to
maintain as a stable rock formation and the objectives of Stability
Control Blasting could not be accomplished.
As the limit between the predetermined stability control zone and
the area that is not to be effected is approached, the weight of
the charge must be successively reduced and in order to get
appropriate breakage and energy utilization, the depth of placement
of these reduced charge weights consequently will have to be
reduced. These relations are controlled by Product Control Blasting
and the Livingston equations and previously discussed. The
significant factor of Stability Control Blasting is that for the
given charge, weight, and depth of placement, the controlling
burden including that for the perimeter holes at the excavation
limit must be such that the limit of extreme rupture does not
extend beyond the stability control zone.
The controlling burden will normally be in the range determined by
.DELTA. = .DELTA..sub.o - 0.10 to .DELTA..sub.o - 0.20. Employing
this method for charge placement, one can thus assure that energy
partitioning does not extend beyond the stability control zone to
the point that the molecular organization of the material is
effected.
Elements of the Mining Method
The mining method is basically the deployment of trajectory and
product control blasting methods in a set procedure. FIG. 24 shows
a mining layout in progress. The ore body is indicated at 10 with
the overburden shown at 12. The overburden bench free face 13 has
been prepared by a trajectory control blast which projected the
overburden that had covered the active mining cells 14 to fill the
previously mined out cells 15. Buildup muck that is not completely
projected to the desired location is hauled to complete the
terracing 16. The terracing 16 is maintained at a slope or grade of
generally 6 percent to the horizontal whereby the terracing may be
used as a ramp for equipment carrying the overburden to terrace
dikes and doms as later described.
As has been previously indicated, the mining layout of this
invention is economical at present basically for mining
applications covering extended areas having the same basic material
characteristics. The layout shown in FIG. 24 could be particularly
adaptable, however, to such deposits as the oil shale deposits in
Utah, Wyoming and Colorado or the Athabaska tar sands deposits of
Canada. The layout illustrated would in these situations represent
an area some 3 to 5 miles wide and 10 to 15 miles long, the cells
being approximately 1 mile long, and the overburden being some 400
feet in depth.
From an original free face preparation, the overburden will be
blasted by trajectory control blasting substantially to the final
resting place of a majority of the overburden and at the same time
expose the ore body. The exposed ore body is then mined by product
control blasting. A number of cells will be planned along the
length of the face and mining will begin in the first of these
cells. The cells will be separated by pillars in order to assure
drainage control, access, reclamation of the surface and water
resources development.
The mining of each cell is conducted using product control
blasting. Since it is necessary to have the ore crushed to a
reasonably small granulometry prior to its introduction to the
milling devices, any portions of the material not sufficiently
broken by the product control blasts will be crushed by crushing
machines in the cells or at the ore preparation plant adjacent the
mills.
Referring now to FIG. 25, an elevation of a mining layout in
progress, pursuant to the mining method of the present invention,
is illustrated. The elevation indicates a face preparation zone 31,
separated from a mining section by a barrier pillar 32 and a number
of barrier pillars 32 are spaced along the length of area through
which the mining process proceeds. The barrier pillars are stable
areas of the same ore body that is being mined and separate the
cellular mining units that are laid out pursuant to the method of
this invention. In the mining layout shown in FIG. 25, a relatively
thin overburden covers the body of ore to be mined. This overburden
may be stripped by conventional mining techniques including heavy
equipment soil removing vehicles or other conventional mining
processes. Once the overburden is stripped from the face
preparation zone, the surface of the mining layout has
substantially the stepped arrangement shown at 33 by limiting the
blast backbreak, the face zone shown being the left face of the
first barrier pillar 32 in the layout illustrated.
In the initial stages of the mining process, test blasts to
determine the characteristics of the material and their response to
the particular explosive to be used are conducted in the face
preparation zone and the proceeding of tests moves to the point
that the face of the mining area is prepared in the manner
illustrated at 33. These test blasts are conducted as previously
described in order to determine the values in the Livingston
equations of N, E, and d.sub.o. Once the overburden covering an ore
body is stripped and the face zone is prepared, the actual mining
process can begin. In the first mining cell, product control
blasting is conducted to break and crush the ore to the desired
granulometry and stability control blasting is used to establish
the barrier pillars in the desired stepped construction. The
blasting in the face zones of each cell is performed to provide
ramps along one face of the barrier pillar, and, referring to FIG.
25, the ramp 34 is shown which provides access to the lower portion
of the cell as the mining progresses downwardly into the ore body.
These ramps are of considerable width and the one illustrated at 34
would be in the nature of some 50 to 80 feet to provide passing
room on the ramp for heavy earth moving vehicles moving up and down
the ramp into the cell and returning from the cell with the ore to
transport same to the milling plant.
Once the first cell 35 is depleted, and usually before completion
of the first cell, the mining begins in the second cell indicated
at 36. The mining of the second cell, similar to the first proceeds
in steps and the first step indicated at 37 is stripped out of the
cell area and transported to the various processing plants.
As the waste from the processed ore accumulates, an area for
disposal must be provided. The waste accumulated from the ore mined
out of the first cell 35 may be disposed of in the open area of the
first cell 35 that has been completed to this point and thus the
terrain re-established for the area in question.
In the oil shale industry, a unique problem not previously faced by
the mining industry must be overcome -- namely, the swelling of the
waste material of an extremely fine nature that must be disposed
of. In the present mining process, the muck that is not projected
to its new position may be used to form terrace ponds designed for
receiving the shale waste materials.
Prior to disposing of the waste in the mined-out cell 35, a water
quality filter bed 38 may be placed on the floor of the mining
cell. This filter bed may be of any standard nature such as
limestone, sand or minute gravel particles, charcoal, etc., and is
designed to control the hydrogen ion concentration of the water and
to filter out the waste and poisonous by-products from the various
ore processing plants so that the water passing through the filter
bed and into streams, rivers, etc., will not contaminate the rivers
and destroy fish and wild life using these water supplies. The
mined-out cell also provides a disposal area for overburden that is
stripped from the adjacent and progressing areas of minings.
In conducting the product control blasting in the mining process
illustrated in FIG. 27, large diameter bores, termed entries and
cross-entries, are drilled through the material. The entries will
be drilled parallel to and in the access entry pillars and the
cross-entries will be drilled through the mining cells
perpendicular to the access entry pillars to provide access of
placement for the charges in the product control blasting.
Alternatively, product control blasting within a given cell or
portion thereof may be done using vertical bore holes as in
conventional open pit mining as shown in FIG. 25.
Irrespective of the depth of the entries and cross-entries, once
the parameters of Livingston's equations are known for the
particular material and type of explosive to be used, the weight of
explosive can be adapted and chosen to gain maximum utilization of
the energy and avoid waste of the explosive charge. In this manner,
it is permissible to drive the entries and cross-entries at
arbitrary depths and the depth of the entries and cross-entries may
be governed by other factors than the weight of the blast to be
employed. These additional factors include the preparation of the
faces of the barrier pillars to provide ramps for access to the
lower part of the mining cell as well as other natural
characteristics of the terrain being mined. Once the depth of the
entries and cross-entries is selected, the value of d.sub.c is
known, and it is merely necessary to equate d.sub.c to d.sub.o and
employ the Livingston equations to determine the appropriate weight
of charge to be placed at the given depth of the cross-entries in
order to gain maximum utilization of the energy of the charge and
thus economize the entire mining process.
As the product control blasting approaches the areas of the barrier
pillars, stability control zones are formed. In effect, stability
control zones are areas in which the charge weight and shape are
changed to eliminate backbreak so as to avoid any detrimental
effect to the barrier pillars. Accordingly, the blast designer will
reduce his charge weight and then choose the value of d.sub.o which
effects maximum utilization of the energy and control backbreak of
the surrounding material in response to the blast. The backbreak of
the material must be controlled so that it will not reach into the
barrier pillars.
The system of entries and cross-entries which extend through the
barrier pillar also provide drainage tunnels for controlling the
flow of accumulated water in the mined-out cells and may be
employed with valving devices to control the flow and supply of
water to the surrounding areas to be explained more fully in
connection with hydro development.
Referring now to FIG. 26, the adaptation of the mining process of
the present invention is illustrated in connection with an ore body
having an overburden of a depth equal at least that of the ore body
and which is greater than what would be called the economical
break-even bench height for trajectory control blasting. This bench
height is where conventional mining equipment would be economically
ineffective as compared to trajectory control blasting to remove
the overburden and expose the underlying ore body. In this
particular arrangement, the face preparation is conducted as
previously described by performing test blasts and, at the same
time, evaluating the parameters of Livingston's equations for the
particular material of the overburden and the explosive employed.
Some crater blast may be originally performed and as the evaluation
of the parameters proceeds, test bench blasts may be conducted in
order to finalize and pin-point the values of the associated
parameters.
In this particular layout, once the parameters of Livingston's
equations are known, the overburden is blasted to a remote area in
order to expose the ore body. Since the depth of the overburden and
the distance .DELTA..sub.bo from the prepared free face of the
overburden can be measured, by using the Livingston's equations and
the parameters previously evaluated the weight and exact placement
of the charge can be determined in order to traject the overburden
to the desired area in response to the explosion of the charge.
In order to determine the exact placement of the overburden in
response to the trajectory control blast, the velocity of the free
face mid-point for the particular material must be calculated per
unit weight of charge. This is done in the test blasting procedure
and is calculated by the formula V = .sqroot.2gh.
When the velocity of the mid-point free face, indicated at 40, is
known, and the angle of trajectory is determined, the exact primary
placement of the overburden can be calculated. The angle of
trajectory is determined by the angle of the free face and the
placement of the charge to control the buffer zone. When the charge
is placed at a distance .DELTA..sub.bo from the mid-point of the
free face and at some distance greater than .DELTA..sub.o from the
surface of the overburden, the angle of trajectory of the free face
mid-point will be the angle complementary to the angle of the free
face itself.
Once the overburden has been blasted by trajectory control blasting
into the mined-out cellular units additional terracing may be
constructed over the filled cellular units from the waste products
of the milling plant in order to conserve the terrain and establish
same at a layout similar to that previous to the mining
process.
FIG. 27 illustrates the use of product control blasting in order to
crush the ore within a particular cell area to a predetermined
granulometry. The product control blasting is conducted in layers
for an ore body of considerable thickness in order to control the
backbreak adjacent the barrier pillars and to avoid disturbing or
weakening the barrier pillars. The layers of ore subjected to the
product control blasting may be of substantially any desired depth
depending on the explosive and type of material so long as the
proper weight of charge is selected by Livingston's equations.
A number of entries and cross-entries are drilled horizontally
through the ore body by heavy drilling equipment and the explosive
charges are placed along the first level or bench 41. The weight of
the charge to be placed in these areas is determined by
calculations using the Livingston's equations. Since the distance
beneath the ore body surface is known and since test blasts have
been conducted in the ore material to determine the parameters of
the Livingston's equations, the formula d.sub.o = .DELTA..sub.o E
.cuberoot.W can be used to calculate the charge weight for maximum
energy utilization as previously discussed. Alternatively, the
charge weight may be selected on the basis of factors involving
.DELTA..sub.b, .DELTA..sub.v, or autogenous blasting. When the
depth of the charge placement is known, the weight of the charge
may be calculated for maximum energy utilization and total energy
partitioning to the material in order to fracture it to a desired
granulometry. Autogenous blasting may also be conducted if a small
screen size of material is needed. The blasting conducted in the
product control zone may be performed basically on consideration of
factors such as desired granulometry of the crushed ore material as
opposed to the additional factors which are considered in the
trajectory control blasting areas.
In the zones adjacent the barrier pillars, the limit of backbreak
in response to a blast must be considered and the depth, weight and
shape of the particular charge must be calculated with respect to
the amount of backbreak that will occur in response to such a
charge and preferably stability control blasting should be used in
order to avoid weakening the barrier pillars or damaging the
entries and cross-entries.
Once a first bench has been completed or is approaching completion
in the blasting and removal of the ore in the product control zone,
it is possible to begin the blasting at the opposite end of the
cell for the second bench or strip to be crushed by the product
control blasting. This progressive movement of mining both
downwardly and in one direction may continue until the limit of the
ore within the cell is reached.
Alternatively, the mining benches within a particular cell may be
subdivided into product control unit blast areas which in effect
permits a step-by-step control of the product blasting as the
crushing of the ore and its removal progress.
The mining method of this invention is unique in the flow pattern
that it provides for the waste materials and valuable minerals
recovered during deployment of the process. The ore processing
plant is located, according to the present mining method, on a
processing plant pillar which is provided at the upper surface of
the ore body and within the confines of the mining field.
By this arrangement, the handling of the material is greatly
facilitated. The overburden is projected by trajectory control
blasting over and down into the openings in the ore body formed by
the previously mined cells. The ore body is then exposed and mined
by product control and stability control blasting with the ramps
formed in the face of the cell for moving the material to the
processing plant either by heavy land moving vehicle or conveyors.
The oil can be recovered from the shale and the spent shale is then
passed further upwardly by pumping slurry to the terrace and pillar
ponds. This flow control is unique in that it permits the orderly
development and reclamation of the land as the mining process
proceeds. None of the prior art mining methods have recognized or
employed this unique flow procedure of the process and have been
faced with a great amount of expense due to the necessity of
hauling the ore to processing plants that are a considerable
distance from the mining field. Moreover, the vast areas of oil
shale deposts in Utah, Wyoming, and Colorado have heretofore been
considered unavailable for commercial exploitation due to the fact
of the swelling of the shale and the unavailability of disposal
areas in which to deposit this tremendous amount of material.
The present method provides areas for depositing this spent shale
by-product and makes economically available a tremendous amount of
oil reserve previously considered by the industry as impossible to
recover.
Hydropower Devleopment and Land Reclamation
Objectives of the reclamation plan for agricultural reuse of the
land are:
1. To provide a soil cover suitable for agricultural
development;
2. To develop a stable land surface free from effects of subsidence
and suited to surface irrigation;
3. To recover for reuse processing plant water discharged with the
trailings;
4. To provide for drainage and control of surface runoff; and
5. To provide a dependable water supply for irrigation.
These objectives are embodied in the reclamation plan of FIGS. 30
and 31.
Land reclamation above plant terrace elevation is accomplished in
three stages -- the first "terrace pond stage," the second "pillar
pond stage," and the third "retreating pond stage." During the
first stage, the reclamation area above plant terrace elevation is
divided symmetrically about the central plant pillar as in FIG. 30,
and in each half, a terrace pond is developed with its long axis
parallel to the pillar. During the second stage, pillar dams and
pillar ponds are constructed in "valleys" between adjacent mining
units. The third stage begins upon completion of cell mining and
provides for reclamation of the T-shaped plan area formed by the
plant pillar and the final row of active mining cells.
Terrace ponds, as indicated in FIG. 30, are formed during the first
stage of reclamation using dykes built of embankment muck. The
dykes are built in a U-shaped pattern starting from the face
barrier pillar end. As cell mining advances, the dykes that form
the sides of the U are lengthened and increased in height until
finally they reach "high terrace" elevation. The high terrace is
the final flat-lying reclamation surface that is to become
agricultural land. It lies at an elevation above the original
ground surface that depends upon the original topography and upon
the overall swell factor of the material as excavated and placed.
The overburden swell factor is reduced somewhat due to deposition
of tailings in the voids of the underlying material and surrounding
dykes. "Low ramp dykes" are constructed across the open end of the
U-shaped enclosure at intervals as cell mining advances and as the
level of tailings pumped into the terrace pond as a
water-slime-sand slurry rises.
An objective of the reclamation is to recover for reuse processing
plant water discharged into the reclamation ponds. The bottom and
sides of the ponds, if constructed of dump-fill embankment muck,
are permeable. The face barrier pillar end of the pond, although
constructed of embankment muck, is zoned, compacted as placed, and
is impermeable. The fast settling sand portion of a slurry pipeline
discharge settles first, and the upper surface of the deposited
tailings slopes away from the pipeline. Water percolating through
the fill carries some sand and slime with it, but the fill acts as
a filter as the water continues downwardly into the mining cells
where it is trapped within the voids of the trajectory muck and
build-up-muck to form an underground water reservoir.
Control of the stratigraphy and elevation of sand-slime deposition
within the terrace ponds is done by lengthening the slurry pipeline
and moving the discharge end as required. The objective is to
prepare a soil cover at final high terrace elevation, and to form a
gently sloping land surface to control drainage runoff. Of course,
the illustration of FIG. 30 cannot clearly show this sloping lay of
the land due to the need for a definite representation of the
various stages of the reclamation process. However, it is to be
understood that FIG. 30 is basically schematic for the purposes of
clarity.
During the second pillar pond stage of land reclamation cell mining
is in progress within the mining unit, an ore processing plant
occupies the V-shaped valley between terrace ponds, and underground
pillar mining may be in progress in an adjacent mining area. The
reclamation layout of FIG. 30 provides for: access to processing
plant, stockpiling of ore at plant terrace elevation, distribution
of tailings either to the ponds or to underground cut-and-fill
slopes, haulage roads between pillar recovery access slots and the
processing plant, and continuing access to all active working
places.
As cell mining advances, a condition eventually is reached where
more space is required for tailings storage than is available in
terrace ponds alone. At that time, the pillar dam nearest the area
where pillar recovery by underground cut-and-fill is being
conducted and the first ramp dyke at the end of the proposed pillar
pond must be completed. Filling of the pillar pond and raising it
to high terrace elevation is done in the same manner as for the
terrace ponds.
Proportions of tailings and overburden to be disposed of depend
upon the thickness of overburden and ore body and vary from place
to place as work progresses. Attention now is directed towards the
flexibility provided by the reclamation layout so as to permit the
required ore waste materials balance to be maintained. The depth of
a terrace pond may be increased to provide additional storage space
for tailings, or decreased to provide additional storage space for
overburden. Similarly, the bottom of a pillar pond may be filled
with overburden to any desired depth if additional overburden
storage space is required, or the bottom of the pond may be built
directly upon the top of the access pillar if maximum tailings
space is required. In addition, the cross section of the dyke that
separates a terrace pond and a pillar pond either may be increased
or may be decreased or the pillar pond stage of reclamation either
may be merged or may be separated timewise to accomodate local
field conditions.
Eventually cell mining advances to its design limit and is
completed in a thickness cell mining unit. At that time, an
alternative step of the method may be chosen, either to fill the
last row of mining cells by trajectory control blasting of
overburden or with tailings confined between boundary limit dykes.
The choice depends to some extent upon the progress of underground
cut-and-fill mining in the adjacent cell mining unit.
In the retreating pond stage of reclamation, tailings deposition
begins at the design mining limit and retreats towards the
processing plant. Concurrently, a pillar dam is constructed on the
highwall side near the processing plant. Doing so permits land
reclamation to be completed and the plant to operate until pillar
mining is completed. Both cell mining and pillar recovery minimg
are completed at the close of the retreating pond stage and the
cell mining unit is left in a condition suitable for the
continuation of hydropower and water resources development.
The mining method previously described is also adaptable to
hydropower development. The valves placed in the openings forming
the entries and cross-entries that remain as stub openings
communicating adjacent cells may be used to control the flow of
water and conserve the water passing into the ground during the
heavy rain periods of the year and then to disperse the water over
a wide area of terrain during the draught seasons. The water
control permitted by this mining method is an integral feature of
the concepts of reclaiming the land under set mining procedure.
Referring to FIGS. 28 and 29 and 30, an exemplary water utilization
diagram is illustrated. During the periods of high water flow,
water will be collected in the lower pools which could be a dammed
river or other convenient water storage facility and can be
selectively pumped to an upper project power plant or to lower pool
plants or passed over a lower spillway. Water passing over the
lower spillway is designed for augmenting the water supply during
low flow periods.
Water pumped to the upper project power plant may be stored in the
open cell water storage area, which in most situations would be a
mine cell that has not been filled by the overburden of trajectory
control blasting or with the spent shale or other by-products of
the milling process. The open cells are maintained specifically for
this feature of hydropower development and provide in effect a lake
in the area of the mining process. The water of the open cell
storage area may be passed through the valve structures connecting
the various cells to the voids of the adjacent cells that have been
filled with the trajectory muck from the trajectory control
blasting.
The mined-out cells form substantially impervious water storage
areas and permit very little seepage of the water between
underground pools, except as controlled by the valving structure.
Thus, by metering the water through the valves interconnecting the
various cells to supply stored water beneath the surface of the
terrain, the mined land is reclaimed for agricultural usage. The
cells adjacent the open cell storage area, while filled with
trajectory muck, may have void storage areas of up to 40 percent of
the volume of the cell. This is due to the loosely packed nature of
the material within the cell.
Drainage collection cells are provided adjacent the voids storage
cells and operate as a reserve collection area into which the water
may be directed during periods of heavy rainfall, melting of snow,
etc.
Water within the open cell storage area may also be passed through
a pump station to the ore processing plant and since the ore
processing plant is located at a process plant pillar on the
surface of the ore body, the use of the water in the open cell
storage area is greatly facilitated. From the ore processing plant
the water is used in a number of ways, one of which is to combine
the water with the spent shale and form a slurry which is pumped
into terrace ponds and pillar ponds which have been previously
formed in the terrain by dykes constructed on the build up
muck.
The water portion of the slurry passing to the ponds flows through
the material of the terrain and into the voids storage cells where
it may pass to the seepage control cells during periods of heavy
rainfall. The water passing from the seepage control cells passes
through filter beds which have been previously placed on the floor
of the mined out cells to filter out impurities which would
accumulate in the water from the ore processing plant. As the water
level of the lower pools recedes, the valves of the drainage
collection cells and seepage control cells may be opened to permit
the water to drain and pass through water quality and drainage
tunnels and water conditioning facilities and then to the lower
pool.
The terms "void storage cells", "drainage storage cells", "seepage
storage cells", etc., are characteristic of the function of the
cells in these physical relations to the open water storage cell as
illustrated in FIG. 28, these cells all being previously mined
cells that have been filled with trajectory muck.
Thus, the mining method which employs the mining cells, as opposed
to the standard procedure of complete ore stripping, greatly
facilitates the land reclamation and water control by the
hydropower development phase of the invention.
Water resources development depends in the method of the invention
upon elevating water from a "lower pool" to open cell reservoirs,
collecting and utilizing underground water that exists within the
ore body and overburden, and upon collecting and utilizing surface
runoff. Normally, the upper pool for a conventional pumped storage
hydropower project is fed by large streams. The method of the
invention introduces a new set of conditions. It is the lower pool
rather than the upper pool that is fed by the stream; and the upper
pool is developed high above the valley floor in workings excavated
during Breakage Process Control Mining.
The hydropower and water resources development plan embodied in the
method of the invention depends upon the layout and geometry of
Breakage Process Control Mining. Although the plan includes a
conventional lower project for pumped storage power generation, the
upper project layout departs from conventional practice, and is
part of the invention.
A conventional upper project (in addition to electric power
distribution facilities and a power plant containing reversible
turbines) receives water from upstream, and consists of a dam, an
upper pool, diversion tunnels, cofferdams and spillway. Employing
the method of the invention, water is received from the lower pool
and a face barrier pillar replaces the dam. A combination of upper
pool units, each unit consisting of an open cell reservoir, a void
storage cell developed during cell mining advance stage, and
completed pillar recovery units, replaces the upper pool of a
pumped storage hydropower plant. Diversion tunnels are replaced by
a system of filter beds, drain holes, and tunnels for seepage and
water quality control. Cofferdams are eliminated; the spillway of
conventional dam construction is replaced by an overflow structure
and segments of trajectory control cross entree tunnels between
adjacent upper pool units.
The hydropower and water resources development plan embodied in the
invention includes the following elements:
1. Diversion of normal stream flow to storage during periods of
high stream flow and return of the stored water to the stream
during periods of low stream flow.
2. Pumping of lower pool water to ore processing plant elevation
with upper project turbines using "off-peak" power, and generating
electricity from upper pool unit water storage during peak
demand.
3. Recovering, conditioning, and reusing ore process-plant
water.
4. Collecting underground water and surface runoff so as to
supplement the normal stream flow and preserve the existing
drainage pattern and water sheds.
5. Developing land for agricultural reuse by depositing in a system
of terrace ponds and pillar ponds a slime fraction of conditioned
tailings.
6. Utilizing the cell structure of Breakage Process Control Mining
to develop an irrigation water supply for the reclaimed
agricultural land.
Before describing the manner in which hydro-power and water
resources development evolves from the mining and reclamation
layout, it is appropriate to consider the significance of the water
resources development plan as it relates to a possible future shale
oil industry in Colorado, Utah, and Wyoming.
Heavy demands have been made by downstream Colorado River water
users upon water, which is in short supply in the arid Western
United States. Due to downstream water commitments, diversion of
water during periods of excess supply and its return to the stream
during periods of inadequate supply are essential to the
development of a shale oil industry. Before a shale oil industry
can be developed the following problems must be solved:
1. Disposal of spent shale
2. Control of subsidence
3. Reclamation of the land surface
4. Efficient conservation of the oil shale and associated nahcolite
and dawsonite.
5. Preservation of existing drainage patterns.
6. Systematic exploitation of a vast area of land that occupies
parts of three states.
The above problems cannot be solved using conventional methods;
but, on the contrary, are solved using the method of the invention.
In addition, water stored underground in the rocks of three natural
basins is utilized; waste canyon land is converted to usable
agricultural land; the region is benefited by a comprehensive water
resources development plan financed by industry and contributed to
society; a growing demand within the region for water and electric
power, otherwise difficult if not impossible to meet is satisfied;
and a comprehensive plan is provided for systematic development of
16,000 or more square miles of land.
A portion of the idealized layout is shown in detail in FIG. 30 and
is based upon the assumption that mining commences at the outcrop
and advances in a series of parallel slices. To apply the idealized
mining and reclamation layout to hydropower and water resources
development, the powerplant must be located as adjacent the outcrop
as shown in FIGS. 28, 30, and 31 to receive water backed up by the
lower pool, and the open cell water storage reservoir must be
located so as to receive water from within voids of the mined-out
cells.
In the course of mining, cell mining units are converted to pillar
recovery units. Interior pillars of a pillar recovery unit are
extracted during the underground cut and fill pillar recovery stage
of mining, leaving the exterior pillars intact as permanent
boundary pillars. The term upper pool unit refers to a complex of
cell mining units and mined-out pillar recovery units
interconnected by cell water passages. The complex is composed of a
series of smaller segments each at one time having been a cell
mining unit. A single, open-cell water reservoir if properly
located at the down-dip corner serves the entire upper pool
unit.
Referring again to FIGS. 28, 30, and 31, the face barrier pillar
and plant processing pillars are permanent pilars between adjacent
upper pool units. Overflow structure and lined spillway tunnels
near the open-cell storage reservoir take the place of the channel
type spillway of the conventional hydro-power construction. They
prevent upper pool units from overflowing the face barrier pillar,
and route excess water to void storage in cells of an adjacent pool
unit in an earlier stage of development. During the earlier stage,
cells increase in number as mining advances, and water is needed to
fill them. Excess capacity for water storage always is available in
seepage control cells.
To be economical, hydro-power units must be of high capacity and
dispersed in accordance with electric power load requirements.
Individual pool units may be served by a separate off-peak pumping
plant at lower pool elevation or by a spillway system fed from a
nearby up-dip hydro-power pumped storage plant. Several upper pool
units can be served by a single hydro-power plant designed to
increase in pumping and generating capacity as additional upper
pool units are completed.
FIG. 28 is idealized and identifies elements of the layout during
an early stage of work. The processing plant pillar is the
permanent pillar that separates adjacent upper pool units. The open
cell water storage reservoir is serviced by the intake channel,
powerhouse, pressure tunnels and penstocks, intake structure, and
processing plant pump station. The adjacent mining unit of the
figure is the first of a series of additional upper pool units
which may or may not be equipped for hydropower generation
depending upon projected needs, water resources functions of all
upper pool units are similar.
As mining advances, each mining cell functions first as a seepage
control cell, next as a drainage collection cell, and finally as a
void storage cell; thus, the volume of underground voids storage
increases with time as cell mining progresses. Prior to pillar
recovery, water is withdrawn from void storage cells of a given
mining unit and is discharged into the lower pool. Following pillar
recovery, the area between pillar recovery boundary pillars, except
the row of seepage control cells alongside the face barrier pillar,
may if desired be reconverted to void storage.
Seepage control cells, except those opposite the open cell water
reservoir, take the place in conventional dam construction of grout
barrier and drainage tunnels. A seepage control cell is a
muck-filled cell containing little water. It may receive seepage
water from adjacent water-filled cells or from the orebody, but
normally is kept dry by pumping or grouting. Seepage control cells
serve to keep water out of active mining cells. In addition to
their location near the active mining cells they are placed to form
a seepage barrier between void storage cells and the face barrier
pillar. Seepage water that collects within a seepage control cell
passes through a water-quality filter bed and is removed through a
complex of drainage holes drilled from seepage control tunnels into
the cell bottom.
Water pumped from the lower pool during the pumped storage
hydro-power pumping cycle is discharged into the open-cell storage
reservoir. Cell water-passage tunnels, driven through pillars
between cells during cell development, route the water into and out
of void storage cells so as to balance the hydro-static head on
adjacent sides of the pillars during pumping and generating. Water
of the upper pool void storage complex equivalent in conventional
practice to the power pool is returned to the lower pool during the
power generating cycle.
During filling of voids of the void storage cell complex, the
hydraulic gradient slopes away from the water level of the open
cell water reservoir depending upon the elevation of water in the
drainage collection cell. The water level rises in a drainage
collection cell as surface run-off and underground water from
overburden and orebody is received. When filled, a drainage
collection cell becomes a void storage cell and supplements the
water supply diverted from streams or diverted to storage from
surface run-off and the spillway system of an adjacent upper pool
unit.
Recovery, reconditioning, and reuse of water are integral parts of
the water resources and hydro-power development plan. Water in
tailings from the processing plant is recovered by downward
percolation from terrace ponds and pillar ponds into void storage
cells. Reconditioning usually is necessary due to requirements of
plant growth for reclaimed agricultural land, oxidation of sulfides
in overburden and ore, or due to reagents added or produced during
processing, and may be accomplished:
1. At the processing plant prior to slurry pipeline pumping,
2. In filter beds at the bottom of mining cells,
3. At a water conditioning plant receiving the flow issuing from
drainage tunnels, and
4. In the open-cell water reservoir as a result of surface areation
or sand and slime settlement from water that is returned
periodically from void storage during the hydro-power generating
cycle.
FIG. 29 is a water utilization diagram which identifies the several
routes for water entering and leaving the system. The diagram if
used in connection with FIGS. 28, 30, and 31 extends the materials
handling system of the method of the invention to include water as
a natural resource.
Attention is directed to the list of objectives of the hydro-power
and water resources development plan--specifically to Item 4, -- as
it relates to the development of a shale oil industry in Colorado,
Wyoming and Utah. The Colorado River drainage system has cut
through the oil shale measures leaving, in most places, the oil
shale beds high above river elevation. Sites for conventional dams
or lower projects are chosen along tributary streams so as to
pre-determine pool limits and to back water up to where diversion
to open-cell water reservoirs is practical. In effect, the design
and layout for hydro-power and water resources development is
determined by, rather than independent of mining requirements.
The method of the invention introduces a new mining technology and
provides a comprehensive development plan and layout in which
mining, land reclamation, hydro-power and water resources
development are integrated to the benefit of society. As mining
progresses, a vast underground water storage system free from
surface evaporation in an arid climate is developed, and the land,
rather than being ravaged and destroyed, is improved to the benefit
of mankind.
APPENDIX ##SPC1##
table 2
---------------------------------------------------------------------------
breakage process equation parameters -- ores and rocks
explo- sive Material Vo/W o E B
__________________________________________________________________________
60% Giant Thin Bedded Gel Magnetite 57.6 .45 4.60 .594 Hi Cap Sandy
Siltstone 47.4 .645 4.20 .63 ANFO Hard Blue Iron Ore 46.7 .65 3.7
.915 60% Giant Gel Granite 44.7 .55 4.15 .622 Hi Cap Arkose
Sandstone 41.7 .59 4.25 .66 60% Giant Gel Magnetite-Taconite 40.4
.48 4.35 .49 ANFO Soft Blue Iron Ore 39.8 .90 3.2 1.20 50% Forcite
Greywacke 37.8 .48 4.0 .59 60% Sandstone Seismo Conglomerate 35.5
.39 5.30 .24 Slurry Iron Formation 34.3 .58 4.26 .44 60% Giant
Magnetite 23.7 .45 4.35 .29 TNT, C2 Navajo Sandstone 22.8 .41 3.3
.63 C3 Unaweep Granite 15.4 .40 3.9 .22
__________________________________________________________________________
TABLE 3
AUTOGENOUS BLASTING DESIGN SUMMARY
CODE 1
Predom Design Particle Criteria Parameters Vol, cu ft .DELTA.
BH=.75N E=3.30 Top 1.0 .DELTA..sub.v = .61 1/2 Col=.24 A=.97 = .79
Sub=.10 B=.47 .DELTA..sub.y = .DELTA..sub.o - .07 V=.375N.sup.3
.DELTA..sub.v trajectory ABC=.375 V/W=13.48 V.sub.o /W=16.9 AC=.798
C=.823 E.sup.3 =35.94
AC= (V/W)/(V.sub.o /W)
CODE 2
Predom Design Particle Criteria Parameters Vol, cu ft .DELTA.
cg+mid-.12 E=3.30 Face 0.10 .DELTA..sub.b = .415 .DELTA..sub.v =BH
A=0.20 V.sub.w =0% .DELTA..sub.v =.75 1/2 Col=.14 B=0.47 V-W.sub.w
= 100% .DELTA..sub.s =.+-. .60 Sub=.14 V=.1275N.sup.3 =.66 Find
.DELTA.b ABC=.1275 =.64 .DELTA..sub.s =.pi. d/6=.60 V/W=4.58 BH=.75
.DELTA..sub.mt =<.DELTA..sub.rc V.sub.o /W= 16.9 ee. Find
.alpha. AC=0.271 .DELTA..sub.cr .ltoreq. .DELTA..sub.v C=1.26
Inclined trajectory
CODE 3
Predom Design Particle Criteria Parameters Vol, cu ft .DELTA.
.DELTA..sub.b =.DELTA..sub.a = .49 E=3.30 Face 0.25 .DELTA..sub.b =
.49 cg below mid A=.25 Top 0.80 .DELTA..sub.s =.+-. .67 1/2
Col=14,.25 B=.47 V.sub.w =40% .DELTA..sub.v =.71 Find .DELTA..sub.v
V=.229N.sup.3 V-W.sub.w =60% f =.22 Find .DELTA..sub.diff ABC=.229
e. .DELTA..sub.mt <.DELTA..sub.rc V/W=8.2 = .76 Steeper
trajectory V.sub.o /W=16.9 =.61 BH=0.75N AC=.4852 .DELTA..sub.inc
=.07 C=1.95 Sub=.10
C=V/N.sup.3 AB
CODE 4
Predom Design Particle Criteria Parameters Vol, cu ft .DELTA..sub.b
= .DELTA..sub.o - 0.10 E = 3.30 Face 0.48 b= .58 cg + mid- 0.11 A =
.44 Top 0.75 v =.70 Col =.14, .25 B = .47 V.sub.w =26% s =.68 Find
.DELTA..sub.v V=.428N.sup.3 V-V.sub.w =74% Sub= .09 .DELTA..sub.s
=.DELTA..sub.o ABC =.428 .alpha.=18 Find Sub V/W= 15.4 f =.12 Find
.DELTA..sub.diff V.sub.o /W= 16.9 BH= 0.75N AC= .9112 Find .alpha.4
.alpha.3 C= 2.07 E.sup.3 =35.94
C.sub.o =V/N.sup.3 B
CODE 5
Predom Design Particle Criteria Parameters Vol, cu ft
.DELTA..sub.rc .DELTA..sub.r E= 3.30 Face .03 v= .61 cg+ mid- .02
A= 0.20 b= .32 .alpha. = 18.degree. B= 0.47 mt= .61 1/2Col= 0.24 V=
.07866N s=.38 Sub= 0.10 ABC= .07866 rc= .53 .DELTA..sub.b <
.DELTA..sub.a V/W= 2.827 t= .58 Find .DELTA..sub.v, .DELTA..sub.b
V.sub.o /W= 16.9 ee. .DELTA..sub.mt= .DELTA..sub.v AC= 0.1673
.DELTA..sub.diff f= .29 Find .DELTA..sub.s C= 0.84
TABLE 4
PRODUCT CONTROL BLASTING DESIGN SUMMARY
.DELTA..sub.b --PRODUCT CONTROL
Predom Design Particle Criteria Parameters Vol, cu ft .DELTA.
.DELTA.-Inc .07 E= 3.30 Face 1.70 .DELTA..sub.b = .73 .DELTA.-Diff
.15 A= .95 Top 3.25 .DELTA..sub.v = .88 .DELTA..sub.V>
.DELTA..sub.T B= .47 V.sub.w =23% =.95 .DELTA..sub.B .DELTA..sub.T
V= .9038N.sup.3 V- V.sub.w = 77% .DELTA..sub.s = .79 .DELTA..sub.B
=.DELTA..sub.bo - .02 ABC= .9038 = 1.11 .DELTA..sub.V
=.DELTA..sub.B +.DELTA..sub.Diff V/W= 32.5 Shape BH= 1.12
.DELTA..sub.mt = .DELTA..sub.f -.05 V.sub.o /W=16.9
1.4.times.1.0.times.0.7 1/2Col= .30.UPSILON.5 2 Find .DELTA..sub.s
AC= 1.92 Sub= .06 .DELTA..sub.o = .68 C= 2.02 =.75 .DELTA..sub.bo
=.75 C.sub.o = 2.67 = 1.15
TABLE 5
.DELTA..sub.v --PRODUCT CONTROL
Predom Design Particle Criteria Parameters Vol, Cu Ft .DELTA.
.DELTA.-Inc. =.07 E= 3.30 Top 3.15 .DELTA..sub.v = .88
.DELTA..sub.v =.DELTA..sub.f =.88 A= 0.37 Face 3.8 .DELTA..sub.b =
1.00 .DELTA..sub.b =.DELTA..sub.ss =1.0 B= 0.47 V.sub.w 21%
.DELTA..sub.s = .82 .DELTA..sub.diff = 0.12 V=1.1242N.sup.3 V-V
.sub.w =79% =1.085 .DELTA..sub.s =1.2 .sub.o ABC=1.1242 =1.12 Find
.DELTA..sub.mt V/W=40.44 Shape BH=1.02 Find Cr/.cuberoot. W V.sub.o
/W=16.9 1.5.times.1.0.times..65 1/2Col=.20 .DELTA..sub.o =.68
AC=2.39 Sub=.06 .DELTA..sub.bo =.75 C=6.48 E.sup.3 =35.94
V/W = E.sup.3 ABC ##SPC2## ##SPC3##
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