cdigitool.library.mcgill.ca/thesisfile61685.pdf · 1102 ,eb stope pillar destress 1902 e stope...
TRANSCRIPT
c:
(
'c" . , J,
'f
\
• f
"
. '
S'rUOlZS or DZSTRISS BLASTlNG AT CAMPBELL UD LAD MIn ..
~ Michael Cullen
Department of Mining and Metallurgical Engineering McGill University,
Montreal.e
--~
,July, "1988
A thesis submitted. to the.Faculty of Grpduate Studies and Research in partial ~ulfillment of the requirements for the D
degree of Master of Engineering.
@ Michael Cullen, ,1988.
'( '~.#... -
..
( , .
Cl ..
. ..
/
•
ABSTRACT
l1any- "mines are currently experiencing sev'ere rockbursting problems as a result of mining to.deeper depths with ever increasing extraction ratios. ~n recent years renewed interest has been given to destress blasting as a means of rockburst 'control. Destressing involves the blast fracturing of rock to low~r the deformation modulus, peak strength, and residual strength~ If the rock is sufficiently fractured the occurrence of rockbursts may be prevented.
Very little information is available on the amount of fracturi~ required for a successful destress blast, and with what blastlng pattern it can be achieved. This thesis attempts to answer these questions and develop guidelines for destress blasting by means of experimental studies of blast induced fractures. Blasting studies were conducted at Campbell Red Lake Mine in Northern Ontario, canada" The experimental results are site specifie; howe~er, the general conclusions are applicable ta fracturing of any confined rock masse W
RESUME
Plusieurs mines ont présentement des problèmes de coups de terrain intenses dus â 1.' exploi tation en profopdeur de mines avec des taux d'extraction élevé S. Réœemment l'intérêt envers le sautage de décompression à été renouve~lé comme moyen de contrôler les coups de terrain. La dédompression implique la fracturation de la roche au moyen de s~utage pour diminuer le module de déformation, la résistance ulfime e~ la résistance résiduelle. si la roche est suffisamment fracturée, les coups de t~rrain peuvent être évites.
,Très peu d'information est disponible à propos du niveau de fractura~ion~nécessaire pour la réussite d'un sautage de décompression, et avec quel modèle la réussite du sautage peut être ré~lisée. Cette thèse tente de répondre à ces questions et donne les directives pour le sautage de décompression par moyen d'études expérimentales de fracturations induites par le sautage. Les études de sautage ont été réalisées la mine Campbell Red Lake, située au Nord de l'ontario, au Canada. L'emplacement des résultats expérimentaux est spécifique; cependant, les concLusions générales peuvent s'appliquer à n'importe quel fracturation, de massif rocheux confiné •
, ,
f.
·r \
\
\
.\ --- --~- --~-~--L\-------------:----------~ ..
AClUlOWLEDGMENT
r' am very grate~ul to Campbell Red Lake Mines Limited for financial support and for providing the opportunity to conduct in
situ tests. Special thanks are given to Tony Makuch, Mike Neumann and Brian Lokhorst t mine engiheers, for their advice, an~ assistance with the in situ testing.
Or. B. Mohanty of C.I.L. Limited is thanked for his assistance
with sorne ~f the laboratory testing of rock ~echanical . properties, as well as many useful discussions on explosives and
blasting.
Assistance with laboratory tésting given by Kerry McNamara and Keith ~tani, of Mc~ill University, is' gratefully',
acknowledged. Many colleagues and staff at MCGiil provided
assistance through discussions on~arious aspects of this work,
for which they are thanked. l am particularly gratefu~ to Dr. M. .Scoble~ my thesis advisor, for providing guidance-during the
course of th!s project.
/'
r ..... .-.1
.r ..
-'
;
!
)
1
1 -6'
/ /
o
/' ,
)
INDBX
1
2 2.1 2.2 2.3
3 3.1 3.2 3.2.1 3.3 3.4 3.4.1 3.4.2 3.4.3 3.5
//4 4.1 4.2 4.3 4.3.1 4.3.2 4.3.3 4.3.4 4.3.5 4.3.6
5 5.1 5.2 5.2.1 5.2.2
6 '5.2.3 5.2.4
5.3 5.4 -
6' 6.1 6.2 6.3 6.4
7 7.1 7.2 7.2.1 7.2.2 7.2.3
! INTRODUCTION Î, CAMPBELL RED LAKE MINE GEOLOGY AND MINE LAYOUT ROCK PROPERTIES AND FIELD STRESSÉS ROCKBURST HISTORY
. DESTRESSING FOR,ROCKBURST CONTRQL MECHANICS OF ROCKBURSTS ' MECHANICS OF DESTRESSING
III •
-
BEHAVIOR OF FRACTURED ROCK I,,;,.~:\;:.,, NUMERICAL MODELING FOR STRESS ANALYSIS ,,'~- -4:,:; THE EXPLOSIVE MECHANISM AND ROCK ERACTURING ... ~'<'I. FRACTURE CREATION '., PREVIOUS STUDIES OF" FRACTURE EXTENSlo'N ~ BLAST HOLE INTERACTION ~ NUMERICAL MODELING FOR FRACTURE ANALY:SrS
REVIEW OF DESTRESS BLASTING PRACTICES INDICATORS'FOR DETERMINING WHEN AND WHERE -TO DESTRESS BENEFITS AND RESULTS OF DESTRESS BLASTING CASE HISTORIES SOUTH AFRICAN GOLD MINES COEUR D'ALENE MINING DISTRICT SUDBURY AREA MINES KERR ADDISON MINE KIRKLAND LAKE MINING CAMP CAMPBELL RED LAKE MINE
PROBEX-l BOREHOLE DILATOMETER -DESCRIPTION OF INSTRUMENT DEFORMATION MODULUS DETERMINATION DILATOMETER CALIBRATION FACTORS DETERMINATION OF HOLE DIAMETER PROBEX-l TESTING CONSIDERATIONS SENSITIVI~Y AND REPEATABILITY OF MODULUS OF DEFORMATION RESULTS IMPRESSION PACKING " SUMMARY AND CONCLUSIONS ON THE PROSEX-1 DlLATOMETER
MICROSEISMIC MONITORING THE MICROSEISMIC WAVE FORM THE CAMPBELL RED LAKE MINE MONITORING SYSTEM MONITORING DESTRESS BLASTING SUMMARY AND· CONCLUSIONS OF MICROSEISMIC MONITORING
FIELD STUDIES OF DESTRESSING NARROW VErN STOPE FACES FRACTURATION AROUND A SINGLE DESTRESS HOLE EFFECTS OF DESTRESSING ON STOPE REACTIONS MICROSEISMIC MONITORING STOPE CONVERGENCE VISUAL INSPECTIONS
i il
-PAGE
l
4'1 4 6
10
i4 14 20 22' 24
- 27 "<.,,31 :-34
36 42
-43 44 45 46 46 48 51 53 -58 66
82 83 85 88 93 94
96 98
I1l6
107 107 110 113 118
119 119 122 124 128 130
1
c
J;
.1t~-ij ;t :J { , "
"'- • ~ ". J-, .. -
:1'" ~ • ) ,
~ - {
',,1 .-
.~.-. " INDEX 0 COlft'INOlIlD
J ~
(7.2.4. ~ 7.2.5
7.3
8 1 8 .• '1
L
8.2 8.2.1 '8.2.2 ~.3 ~.4
9 9.1 9.2 9.3 9.4
.,-10
, J '. "
1
~ .. ~ . \:\ , OTHER'INSTRUME~JAT~~N DISCONTINUITY SURVEY, SUMMARY AND CONCLUSIONS STUDIES .
; .....,..
OF *REAS,!,' FACE' ~STRE~SINS ,.'
BLAST FRACTURIN~ STUDIES ~FRACTURATION STUD)ES IN THE
PILLAR FRACTURATION STUDIES. IN THE ~OCK,CHARACTERIZATIqN BLAST FRACTURA'.j.fIqN . BLAST VIBRATIO~ STUDIES· SUMMARY ,NND CONCLU,SION$ OK
DESTRESS BLAST DESIGN
\ .
" "-
\~67.0 'w '~T?PE é~~~J . . 1470 STOPE ACCESS DRIFT
, BLAST FRACTURING STUDIES
BLAST HOLE SPACING AND GEOMETRY EXPLOSIVE SELECTION . OTHER BLAST DESIGN CONSIDERATIONS ,-~NUMERICAL MODELING FOR DESTRES5- BLAST DESIGN . .
CONCLUSIONS
APPENDICES
REFE~NCES ,"
1
.\ [.
Il l" , . "
"JI
. .,
, '" .. ,
"
\
1 --~
..
. ,
PAGE , . . 134' .. 135
138
140
140 î47 1'48 151
_ '162 165
1 16-7 171 173
'174 ,
1'76
, 179
196
, Il.
1 Il 11
P il Il
iv
1
,
- "
, . '
..
..
•
:iI, . l'" §. 4 , '-
.->;-
. .{/' 4
L'IST QF FIGURES , . ~ .....
F~~~ -.--FIGURE
" ' ... FIGURE , -FIGURE
. \
2-1: 2-2: 2-3: 3-1:
, FIGURE' 3-2: FIGURE 3-4:
--- , . FIGURE:~" -3-4: FIGURE 3-5:
FIGOfΠFIGURE
3-6: 4-1:
Mine Loc,at ion Mqp f
Beological Plan" 1400 Level JIn Situ Stress 1400 Levei Reactions' To Loading In A Stiff And Soft
, Tej:)ting Machine ' , , - The Init.:i"al· Concept '6f Destressing . . t Normalized -Dynamic Stresses: Step Pulse An'd_
, Wave Train . Zones Surrounding Blast Hole' Four Cases Of Initiation Delay Between Charges 'a' And 'b'
" FIGURE ~21-2: FtGURE,:~ 4-3:
Three Dimensional Fracture Development Stope Pillar Destress Blast Pattern Coeur D'Alene Preconditioning Stress Blast patte~
. l'
FIGUtŒ " 4-4: FIGURE 4-5: FIGURE 4-6: FIGURE ·4-7:
F~GURE 4-8: ' fit
FIGURE 4-9: 4-10: 4-11: 4-12:
FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE F.IGURE FIGURE FIGURE FIGURE .)fIGURE FIGURE FIGURE FIGURE FIGURE
FIGURE FIGURE
- 4-13: 4-14: 4-15: 4-16: 4-17 : 4-18: 4-19: 4-20: 4-21 : 4-22: 4-23: 4-24: 5-1: 5-2: 5-3: 5-4:
'5-5: 5-6: 5-7 : 5-8:
5-9: 5-10:
Stope Pillar Destress Blast Falcon ridge Development Reading Destressing A INCO, Pa~t
~. ." Development Headlng DestreqSlng at INCO . Destress p'attern St ope Backs Kerr Addison Mine Destress Pattern For Raising Up T~ Level, Kerr Addison Mine .
. Destress Pattern For Squarré S~" Stopii1g Kerr Addison Mine Destress Pattern For Pillar Recover~. Destress Pattern For Stope Approach!ng Fault Destress Pattern For Approaching Stopes Destress Pattern For Raises Long Section Macassa Mine Stope Crown pillar Destress Macassa Mine Shaft Sinking Destress Pattern Macassa Mine 1102 ,EB Stope Pillar Destress 1902 E Stope Pillar Destress Long Section Al Zone 1604'EW Stope Pillar Instrumentation 1604 EW Stope pillar Destress Thin Section Photograph Of Microfractures SEM Photograph Of Microfractures -1421 Drift Box Hole Pillar Destress Box Hole Pillar Destress CRLM r
Schematic Diagram O( Probex-1 Wireline Attachment. Probex-1 At Underground Test Site Typical Pressure Versus Volume Test Data Idealized Pressure - Volume Deformation Curves Tracin9 Impressions From Dilatometer Fr~cture Impression Maps Inclined Discontinuity Plane Cuttin9 Drill
" Hale Impression Map From Drill Hole-Equal Angle Stèreograph With Example Solution
v
\
PAGE
5 7
12 -
17 21
28 35
38 41
49 49 52 52 52 54
56
57 59 59 60 60 63 64 67 69 70 71 73 74 77 77 7-é ll
80 64..-86 86 91 91
100 101
102 102 104
c
(ilC
J. -'
LIST or rIGURES CONTlNOBD , ,
. FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE
... FIGURE
FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE
FIGURE FIGURE FIGURE
FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE FIGURE .
FIGURE
6-1: 6-2 : 6-3: 6-4: 7-1: 7-2: 7-3: 7-4:
7-4B: 7-5: 7-6: 7-7: 7-8 :-7-9: 7-10: 7-11: 7-12: 8-1: 8-2 : 8-3:
8-4: 8-5: 8-6:
8-7 :' 8-8: 8-9: 8-10A: 8-10B: 8-11 : 8-12: 8-13: 9-1:
9-2: ,
, Microseismic Event ,-Micr6seismic Monitoring~System Seismic Activity Versus Load Seismiç Survey Layout And Resultjl. Blast Hole Eldngation Blast Hèle Elongation paint Injection Trial 1704 EW~tope, Test Location And Instrumentation ~ Test Blast Sequence Typical Blas~ Rattern Microseismic Activity After Blasting Stope Clos ure ' Stress Fractures Before Scaling Stress Fractures After Scaling Blast B3 Destress Hole Blast C2 Destress' Hole, Discontinuity Map 15 Level Plan. 1p70 W 'A' Stopè
~ , ,
Core Recov~red From 1670 W 'A' Stope' Crown Pillar St reps Strain Curve CRLM Andesite Horizontal Versus Vertical Strain Blast Hole And Diamond Drill Hole Locat~ons 1470 Drift r, D
1470 Drift Destressing Results Hole 3 1470 Drift Destressing Resùlts Hole 4 Core From 1470 Drift Hole 4 1470 Drift Fractures - Impression Packer 1470 Drift Fractures - Drill Core Fractures Versus Blast Hole Spacing Modulus Versus Blast Hole Spacing Modulus Versus Fracture Density Relationship Between Fracture Density And Blast Hole Spacing Relationshlp Between Deformation.Modulus And Fracture Density
,
n
'1 ,
..
vi
PAGE
109 11.2 115 117 121 121 123
125 126 127 129
·129 131 131 133 133 137 141 143
145 146 146
152 154 155 156 158 159 160 160 161
110 •
,170
.\
.. .. Il
LIS'!' 01' 1ABLBS .~
PAGE'
~ -_ 9, TABLE 2-1 :
, TABLE Û2 - 2 : .
Rock Varieties: Mine Terminology ~d Actuâl Rock Types 8 Mine Levels And Elevations 9
TABLE 2-3: TABLE 2-4 : T,ASLE 4-1:
Rock Properties _ 9 Results Of In Situ 'Stress Measurements .. 11
1 Summary Of Results Of Destress Trials 47 TABLE 4-2 :' Stress Ch~nge And Closure 1604 E Pillar ' 75' TABLE 5-1: Confined Calibration Results 92 TABLE 7-1: TABLE '.8-1 : TABLE ,81-2 :
Discontinuity Ori~~tation 136 1670 W 'A' Stope Crown pVllar Test Resulta J 144 1470 Drift Rock Properties 150
TABLE' 8:"3: Results Of Blasting Tests- 164
r
~,
" , ,
( . <1
'0 ""
--" .
=
. . ,\
.,..
o •
vii
/
"
C'
... f ,
" 1. INTRODUCTION
The past fort y years has seen very few new ideas for the control ::-r-
of rockbursts with the exce~tion of\the destres~ blasting méthode Review of the literature reveals that the early mine engineer~ were familiar with the i~eas of convergence control and-;ine
, '1),. J
sequencing as preventative measures foY,rockbursts. However, even with modern techniqu~s of convergence control and mine ~equencing, rockbursts continue to occur.
01
,. _ oestressing inv'ol ves al ter-ing the rock mass mechanical properties
---by.creating blast induced fractures. The method has been met with both,skepticism and support. There is limited documented
accounts of the methods success, although it is believed that many, undoccumented, succe~sfull blasts have taken place. Resistançe to the technique js likely the r~sult of unsuccessful
r blasts and limited experiencei many engineers believe'it to be unproven technology. It is the belief of this author that man y unsuccess~ul destress blasts are due t~ poor planning and execution, resulting in the rock mass being either under or over
./ . . fractured. Another drawback of the technique is that once the ground has been destressed, it is not possible to know what the outcome might'have been had destressing not been occurr~d.
Tne preyenti9n_ of ro~kbursts by d~stressing was first practiced in the deep South African gold mines. Initial results in South Africa were very good, however, as mining continued to greâter
c'
, depths and greater extraction ratios,4bursting persisted. In recent years, with more sophisticated design and evaluation
~
tools, destressing has been proven to be an effective method'of rockburst prevention. It is recognized that a greàter , understanding of the mechanics of de~tress~~ and the design destress blasts are still necessary.) _
,of
l'
The term destressing is somewhat a misnomer, a more suitab~e term ~
o. would be rock mass conditioning. Because of its general acceptance \ . within the mining community, the terms destress blasting and
o
..
destressing will be used throughout this thesis. There are a r
number of o?se~ved and postulated phenomena contributing" to/' _~i~
successfu~ destress blast application. ;
~ a) A' reduction in st-i-ffness due to fracturat'i\;~'. i " • 1
1
b) A. reçiuction ~n }~e/ak, a~Gi residual strength ,d~~. oto f'racturation. c) Contr.olled e'n1'gy dissipation along frac~uFe stirt~ces. d} A reduction in potential energy due to closure. l, e) Non' _violent fai.lure as a result of transition to J.seudC?-
plastic behavior. f) Triggering a controlled rockburst with the blast.
A~l these aspects are covered in general discusss~ons, however, only the reduction in stiffness due to fracturing is\investigated .. expe-rimentally.
v
The term destressing is also used for the tèchnique of degassingc for rockburst prevention in coal rnes, and 'for triggering co~rolled rockbursts,by the infection of wat~ ante joint and faûlt surfaces. These two method~- are not discussed in this paper. ( 1
Campbell Red Lake Mine (CRLM) is a gold mige lbcated in Northern Ontàrio. Gol~ is being ~ined from steeply diPrping, nar~ow vein stopes, by overhand cut and fill tnethods·. The host rocks are andesites and altered chloritic schist. The mine has a history
'of rockbursting starting in the early 1960's. 1 Rockbursta with .1
local Richter-magnitudes up to 3.2-have been recorded. A number
of destress blasts,~ave been conpucted at the mine tith varying . degrees of success. The philosophy at the mine -is ciuse -destressing, for rockburst control in pillars,-as the method of l~st resort. ~his is due to both the high coat of destress blaat!ng and concern over variable reaults. It should be noted that the cost of destressing is below that of placing the required strength an~ modulus backfill necessary for rockburst prevention. Destressing stope breast faces ls carried out routinely at the mine as the need arises. Management and
2
c
\
(
c
. .
c
J
t
engineering, at the mine, recognized that the destress blast designs, which had developed on a trial and error basis, were in need of improvement. Verification of the methods effectiveness was also .desired. (
1 This thèsis is divided into eight sections starting with a;review of mining and geology at.CRLM and a hlstory ot rockbursting at the mine. Chapter 3 presents the meëhanics of rocfbursting and destressing as well as the theory of blast inducedlfractures. An'extensive literatùre review of des~ress blasting was carried out: this is discussed in Chapter 6. A detaile4 account of_~ .heavily monitored pillar destress bTast at CRLM la included in Chapter 4, along with an analysis of drill core, recovered from . the destressed pillar: this analysis was conducted by this author.Microseismic monitoring and in situ deformation testing
, . were used as the principle tools for the studies/of destress blasting carried out during this project. The borehole dilatometer used for determination of in situ deformation modulus
"was a recently developed ins,trument. . techniques needed to be developed, as
instruments validity. Chapt ers 5 and
As such, operating . weIl as verification of the 6 discuss the two , - ,
moni'toring and testing methods. Details of operating' techniques developed and instrument limitations are given. Chapters ~ and 8 report on the two aspects of destress blasting studied. The tirst study evalua~ed single hole destress blasting ofa stope
-=breast face. An investigation of fracturation around Lthe hole as weIl as a study of the overall effectiveness of the technique,was
carried out~ The second study dealt with blast fracturation. «
Investigations of fracture extension be:twee~blast holes' was carried out as well as monitoring of peak particle velocity.'
/ -
Chapter developes guidelines for destress ~last design at CRLM based on the experimental results and literature review. Chapter 10 presents conclusions and recommendations of this study.
3
.. \
I~
./
•
- ,
. C
• ,
2. CAMPBELL RED LA1ΠMID: OVERVIEW
1
CampbeYl Red Lake Mines Limited is a member of the Placer Dome
group. The m~fe is loc~ted in Balmertown within the Red Lake mining district of Northwestern Ontario, Figure' 2-1. Gold with trace silver is mined from thin, steeply dipping ore veins. Mining cornrnenced in 1948 at a rate of 300 t per day. Current
j
production is 1100 t per day .
. Original prOduction was all from shrinkage stoping. Overhand cut and fill stoping is now used almost exclusively below a depth of 200 m. Classified mill tailings, supplemented with sand and -
crushed waste rock, are used for backfill. Cement is added to
sill plugs and stope working surfaces. , Currently, mining is to a - " -depth of 900 m. Access to the mine i5 by a 4 compartment shaft
extending to a depth of 1300 m. It is expected that mining will continue to greater depths in the future. Extraction ratios vary
j
thr~ughout the mine and can be as high as 85 percent in sorne ,areas. 'Hand held,and electric equipment are used in stoping.
,> , The mine has a history of extensive rockbursting which is
ù
generally attributed to the mining method (shrinkage stoping) and ~ .
mine sequencing used in the pasto Geo10gy also has a significant
influence on the occurrence of.~ockbursts. The mine consista of . 5 major ore zones in a comp1ex geologica1 environment. This chapt~r presents an overview of the geolpgy and rockburst history
.;.
of th~e. Details of the mine 1ayout are,given for reference
in subsequent chapters .
2.1
, The mine is ocated in the Red Lake Greenstone Belt which
includes m tavolcan~c and sedimentary rocks. The greenstone belt is part of the B~rch-Uchi lake subprovince,. which in turn 18 part of the Archean Superior Province, Rigg (1980).
4
(
c
,-----
\
f
Red Lake Min:
i n Map 2-1: Mine l~c_at 0 Figure
\
5
,1
'\ -
Ont ario
/
..•. , , f
•
•
. Hydrothermal intrusions and tectonic deformation have significantly altered the rockmass such that identification of original petrology and deformation seque~ce is not possible. Figure 2-2 shows a simplified geological plan of the 14th level. Rigg (1980) 'suggests that the rock classifications used by the mine are ov~~ simplified. Table 2-1 presents the ~esults of his findings.
The deposit is hydrothermal. Multi~phase intrusions have been noted by Rigg c (1980), although not aIl havé been gold bearing. Subsequent remobilization hasrurther co~plicated the mineralization history. The chloritized schist is believed to be an ultra mafic intrusive, Church (1985). This intrusive ljkely caused deformation in the andesite allowing for inflow of hydrothermal solutions. ,
Ore zones occur in a number of geolog~cal settings but are always
~ubparalle: to cleavage and major structures such as andesite -altered rock contact. The mine:Ls comprised of 5 maiQ. ore zo~s, F, F2, .~, L, and G, as weIl as a number of secondary
zones, S, 0, Al, NL, Figure 2-2. The ore veins are of two distinct types; quartz carbonate fracture filled and replacement type.' Quartz carbonate veins are from 0.25 to 1 m in width,
replacement veins are 0.6 to 9 m wide. The host rocks are andes!te and a chlorite schist. Many ore veins occur along the contact of the two. Rockbursting is only a problem in the stiffer andesitic rock. The less stiff chloritic rock may
, -influence whether failure is violent or non violent.
The mine is divided into 27 levels at approximately 45 m , intervals. Table 2-2 lists level numbers and d~pth below
surface .
6
l'
':
•
CI) &al &al t- ~ ...J a: &al u 0 Q z ~ r :x:
"- 0 0
0: ci N
ci ...J W CI)
Cl ...J 0 :::» W ~ \&.1 0
Z Z ID fi) a:: &al Q. W &LI 0
&al W ::1 Cl !:i -' a:: a: oc( z 0 0 0 oc( oc( CI)
lI1~IIID
Figure 2-2: Geoiogical Plan 1400 Level::
c '/
7
•
1'L' 41
•
'l'able 2-1: Mine 'l'erminology and Actual Rock Type •
Mine Terminology
Mdesite
Diorite
Rhyolite
--Altered roçk or Chloritic schist '
Lamprophyre dyke
• J l,
Rock type
P,redominantly basaIt rhyolitic andesite and variolitic andesite'~ith minor i~erflow sedi~ents.
,-
Two massive fractitionated gabbroie bodie~, possibly extrusive. '
Rhyolite
Strongly altered mafiç to ultramafic 'in'trusive rocks, basaIt, andesite, rhyolite and~sediments.
Mafie dyke' , •
after Riggs (1980)
(
, ..
\
..
8
,~
, .
=-."~-
C
\
C
'l'able 2-2.: Mine Level. and Blevations
'" ~~~~l Elevation (m) L~~~l
b~lQ1t :iil.u;:fa.~~ =
.. 1 53 ' ' 14 ,
2 91 \ 15
3 130 16
4 168 17
5 213 ,18 --~,-
6 259 19
7 305 - 20
B 351 -- -=-21
9 396 22
10 442 23
Il 488 24
12 533 - 25 " 13 579 - " '26
.; 27 '"
'-~
'l'able 2-3: R.oc" Propertiea
Rock no. uni axial Young's type samples strength modulus
(MPa) (GPa)
Andesite+ 8 172 82
Altered rock * 2 116 57.5
Diorite* 7 156 88.7
Ryolites * 5 235 92.5
-
, .. "
Elevation (m)
belQw surface
625
670 716 762
808
853
899
944
991 1036
1082
1128) 1173 \ -
1219
poisson's tensile ratio __ strengt
(MPa)
Cl.2 15.6
~ .15
0.26
0.19
* values from Michigan Technological University, Rock Mechanics Research Facility
+ values from McGill University Rock Mechanics Laboratory
9
-
!?
-J 2.2 ROCK PROPERTIES AND FIELD STRESSBS
Mechanical properties of the various rock ,types are listed in
Table 2-3. The values are averages from rock samples recovered from throughout the mine.
,~ 'A,,"number of overcoring type stress measurements were conducted by --
CANMET at the mine using triaxial CSIR cells. Table 2-4
summarizes the results and compares them to expected ~alues based
on.calculated stress gradients in the Canadian Shield of Northern - Ontario. Figure 2-3 shows the location and direction of
measurements taken on-the 14th level of the mine. As ls common
in the Canadian Shield, the major principal stress is horizontal,
striking NE-SW. It is almost double the vertical stress, which
is slightly greater than that expected from overburden lo~ding
alone. \)
2.3 ROCKBURS:' BJSTORY
Campbell Red Mines has had a long history of rockbursts. The
first unlocated rockbursts occurred in the vicinity of the mine
before 1961, Morrison (1961). The first located bursts occurred
in remnant boxhole pillars. They were attributed to the removal . ~ ~ ~ ~
of sill and boxhole pillars in shrinkage'stopes. Later, as ,
extraction ratios increased, bursting was triggered by pulling
ore. from shrinkage stopes and mining in their immediate vicinity,
Hedley et al. (1985). Most of the bursts occurred in the boxhole
and sill pillars and around the intersection of raises with
stopes and drifts. Si Il removai iri shrinkage arèas above the
lOth Ievel has been discontinued. As weIl as pillar bursts,
small, strairt energy bursts have occurred regularly around the
periphery of underground openings. A description of the
different burst types is given in Chapter 3.
In 1983 a series of major rockbursts occurred in the 'F' zone.
This is a shrinkage area with a very high extraction ratio.
10
.... 1» r::r a> 1\)
1 .110>
:0 a> fi)
c -fi) ... 0 ... -:l
Ch -C
Ch -.., a> (/)
(/)
3: l1li a> ... 1» - (/) c»' C .. .., > a> .. 3 l1li CI> :;, :l Ul -... (/)
/t) 0) 0)
?,
~- ~ ~-~
"-
.p
Depth Field Stresses (in MPa), CRL Mine Northern Below Test No. 01 02 a' Ov oRa oHa/av Ontario Data*
Surf~ce Brg./Dip Brg./Dip Brg./Dip % Diff. X % Diff. % Diff. (~Pa)
625 m CRL Tl-2 . 23 lS 10 14 19 1.4 ov=16, oHa-33 . (14 L) CRL Tl-3 060°/40° 214°/33° 134°/33° 12% 42% 6% oHa/ova1.5
combined \
')
.,.
990 m 3 test 53 24 12 25 38 1.5 ,ov=26, 0Ha-44
(22 L) combined 087°/16° 322°/57° 185°/26° 4% 13% 7% oHa/ov-1•4
1220 m CRL T3-2 70 41 31 34 55 1.6 ov~32J 0Ha-47
(27 L) CRL T3-3 230°/10° 020° /76° 140% 7° 6% 17% 14% oRa/"v·,1.4 combined - -_._-_ .... _- _._.~---
x Differences of vertical (ov), averag~ horizontal (ORa) stresses and stress ratio (ORa/Ov) with Northern Ontario data.
* Northern Ontario data, obtained frQm stress gradients in the Canadian Shield
Ov • 0.027 MPa/m (obtained from overburden weight).
ORa • 9.86 MPa + 0.0311 MPa/m (0-900 m depth below sùrface). ,
OHa • 33.41 MPa + 0.0111 MPa/m (900-2200 m depth below surface).
OHa/qv - 251:68 + 1.14 depth m
p'
\..,
~
,.
- 1
1
\
\
'-,
"
o
o trl
1
--'\
W
~-.J_ W g
.J > -0",' 10
W " -1 -
W .J
o W
sl=major-princjpal stre8~ s2=intermediate stress .J ~ s3=minor principal strestd
~ (/)
see Table 2-••
-- .
al a... ~ « u
Figure 2-3: ln Situ Stress 14,00 Level
after Ma uch 1186
12
...
---,/
.'
"
c •
L9cal Richter magnituêles up to 3.1 were --recorded. The area is now c10sed off, leaving a large quantity of broken ore within the stopes. A detailed description-and analysis of this bursting activity is presented by Hedley et al. (1985). Rockbursts still o~cur in the zone due to the slow decay of supports and stress transfer from other areas.
More recently, rockhursts have occurred in crown and sill pillars'of eut and filt stopes. 'Pillars i~ stressed ground are prone to bursting when they reach a critical size, dèpendent upon' stress
'leveis and geology': Under present mining conditions, 10 m is the , '
critical siz~. A fault slip type burst has also been identified in a eut and fill stope, . Neumann (1986).
As weIl as the loss of ore and cleanup costs, rockbursts present . a major threat to life. Fortunately Campbell Red Lake Mipes.has
had no fatalities as a resuit of rockbursts. In order ko quickly and accurately located rockbursts and other sei smic activity the
imine has installed a 64 channel microseismic monitoring system. ,
A description of the system and its uses as weIl as a discussion of microseismic emisaions ls presented in Chapter 6. High energy
microseismic events have been recorded in 16 out of 50 current~y active stopes. Low energy events are recorded from most stopesJ particularly after blasting.
. "
" -'1
13
)
o
3. DBSTRESSING FOR' ROClœ'ORST OON'l'ROL
The past fo~ty years have seen very few'new ideas for the'control~ of rockbursts with the exce~tion of the destressing rnethod • .. Review of the early literature, ~obson (1946), reveals that the early mine engineers were familiar with the ideas of ,'closure
o contr91 and mini~g sequence for rocthurst con'trol~_. !l'he effects -~= 1 1
of time, 'geology and depth were also recognized. Even tnough modern analysis, and techniques tor contro11ing convergence, are much more sophisticated, and minlng sequence can be planned to minimiie high stress situations, rockbursts still occur.
"
This'chapter discusses sorne of the theoretical considerations of destress blasting. As an introduction ~o the problern a review of rockburst mechanics is presented. This is fo11ow~by the mechanics of destressing. Destressing involves the use of
explosives ~or blast fr~cfuring. An extensive r~view of the explosive mechanism and its relationship to fracturing is 4
presented along with discussions on numerical modiling for stress analysis and fracture prediction. Where applica~le, comparisons are rnad~ between theory and practice at Camp~ell Red Lake Mine , (CRLM) •
. ' 3.1 MECBANICS or ROCKBURSTS
The ~echanics of a rèckburst are not completely understood, although, a numbei of hypotheses haye been présented. The commonly accepted causes are related to the disturbance of the gràvitationa~ and tectonic forces due to excavation of underground openings. The creation of excavations causes stress
and strain energy concentrations about the excavation. If the ;
stress concentrations exceed the strength of the rock, fallure "
will occur, causing closure. A rockburst i8 the result of violent failure. It is interesting to note that rockbursts have occurred in surface quarries although these are not considered in
this work.
14
c
c
, If the rockmass is assumed to behave in a purel~_elastic manner with no fracturing, formulation of energy b~lance~quations to describe reactions of the rock mass to mining i5 simple. The formulation i5 covered by many authors including Hedley (1987)
\
and Salomon (1984). The general form of the equation follows:
.,.
.,
where Wt = change in potential energy resulting from
Um , ' Uc
Ws Wr
ex"cavation ,.J
= stored strain energy in excavated rock = increa~e in stored strain energy in surrounding
4-
rock = energy absorbed in support deformation = excess energy released
'/" "'. ..
The sei smic energy-' Wk is responsible for damage to the mine; and is a resul t' of c~osure of the excavation if pJ:1_r~ly --elastic • ..
Assuming elastic behavior the se1smic energy can be shown to be:
..
. A number cvenergy relationships'can be d~veloped, from which the
foll~wing statements can be ma~~1 af~er Hedley (1987) •
.. a) Without support, aIl the energy components can be expressed in terms of two parameters Um and Uà (the increase ~n stored strain energy if the stress increase ha occurred on an unstressed specimen). It is relatively easy to calculate these parameters using numerical techniques. b) When mining takes place in very small steps the process is stable and no seismic energy is released. In sorne cases this is contradictory, since many mines employ incremental methods but still experience rockbursts. However, this means that sorne other sources qf energy are being liberated due to non-elastic 90ndition"s, either fracturing of a pillar or slippage along a fault. c) The change in potential energy, Wt is the driving forcé oehind the energy components. If it can be reduced, the other energy components are correspondingly reduced. d) Su~port, su ch as backfill, has two beneficial effects. It
15
•
~ >
7 o
will reduce th~change in potential energy by reducing volumétrie cl9sure in the excavations and by absorbing energy, less energy is available to be released as seismic energy.
In reality, the energy released by ruckbursts cannot pe'desc~ibep by elastic reactions, as fracturing and other non elastic behavior occurs.
The seismic energy radiated out is a result of tpe oscillations of the rockmass a~out excavation boundaries , as failure occurs.
,)
Damped oscillations result as the~ock mass closes in on excavations. The amount of seismic energy released is prop~rtional to the size of the failure. ,
J For violent failure to occur, the stiffness of the mine structure must exceed that of the mine loading system. The stored energy of the-mine loàding system ~s imparted into the fai~ing structure 'causing violent failure. An analogy is drawn to a core specimen being loaded in a testing machine. The amount of strain energy
, stored by an object is inversely prop~r:tional to its-stiffness, stiff rnaterials store less strain energy.
~~-
w = stored=-strain energy
fi = stress \ a =_cross-sectional area
l = length-K = stiffness = aE/I E = modulus of deformation
When the stress imposed by the testing machine exceeds the st~ength--?f "the specimen, the relative stiffness oi- each determines whether violent or nonviolent failure occurs. 3~1 shows the stress strain reactions for the two possiçle
J
\
16
Figure
c:a8es.~
•
(
. ,
c
. (
Figure 3-1: Reactions To L,oadlng ln A Stlff
Soft Te.tlng Mf chine' -
F
AvoJloble stroin energy, ' ,
DISPLACEMENT CASE A: U nstable: violent "",foilure
17
u
F
w -4 u
cr ~
DISPLACEMENlf 'CASE B: Stable, nonvio.1ent
after B la'ke 1972)
...
\, . ,
r:: case (a), the specimen stiffness (Ks) is greater than the~· loading mac~ine stiffness (KI). When the load exceeds the specimen st~ength, violent failure will occur. The additional .. (\ -stored strain energy in the loading machine continues to load che specimen as it recoils. For case (b), Kl is greater than Ks. Nonviolent fa~lure will occur as the stored strain energy in the loading machine is not suffiëiênt to further load the sample. Tc cause greater damage, additional l~ad must be applied by the testing machine. ~
Whether or not a mine pillar fail~ vlolently ls a bit more ,-complicated than discussed above. It ls dependent on the post failure 'stiffness of the pillar compared to the stiffness Of the
o
entiFe loading system., The size, number and location of the pillars influences the loading system stiffness.
A rockburst causes instantaneous fracture and dilation of the ground. This causes shocK waves to be propagated through the rock as weIl as d~splacement of rock towards mine,openings. The
shock waves ~ay cause rock to 'burst' off excavatio~ wailS and cause other distortion and movement. An opening close tO a
rockburst may be fille~by ~displaced rock; this is usually in a
violent manner.
Salomon (1983) suggests that the follow~ng conditions are n~cessary -~
for a seismic event, hence rockburst, to occur.
i) A region in the rock mass must be on the brink of unstable equilibrium due to:
a) the presence of an appropriately loaded pre-existing geological weakness, b) changing stress driving a volume of rock towards sudden failure (can be an increase or decrease depending on type of rock burst), ~- ~' c) pillar support systems approach,ing a state in which ' the' unstable collapse is imminent.
i~) Som induced stress changes, regardless of magnitude, must be suffici ntly large to trigger the in~t~bility. iii) sudden stress change of sizablel amplitude must take place at the locus tability to initiate the propagation of seismic iv) Substantial energy must be stored in the loaqing
, 18
c
•
c
system, as discussed above. The orïgin of this stored' strain energy is work done by:
a) gravitational forces, t b) tectonic forces, c) stresses induced by mirring.
The three categories, listed in Section (i) of Salomon's ~ criteria, are commonly referred to as fault-siip, strain and
p~llar bursts respect~~ely. Fault-slip t~pe bursts are identical t~ earthquakes and, caf}" pe remote from mine workings, whereas, the
other two occur withln the mine workings. Strain bursts occur near working faces antl are attributed ~o the lowering of the rbck strength ·due to removal of confinement. Pillar bursts are the
result of increas~g stresses beyond the strength of an-isolated
rock structure. The increase in stress 1-s due to stress transfer
from adjacent areas or reduction in pillar size.
A significant stress\change and large quantity of energy is necessary to initiate~ismic waves, originally the energy in
~
rockbursts was attributed only to the stored strain energy. It is noW apparent that changes in the potential energy contribute
significantly to both strain and pillar type bursts. The
different burst types have different seismic effieiencies, this
~~~e percentage of total energy released, released as seismic energy. Much energy is used in the creation of fractures, slip
movement and the generation of heat: Typically, strain bursts
have a 30 to 60 percent seismic efficiency, pillar bursts a 70 to
90 percen~ efficiency, and fault slip bursts alto 10 percent
seismic efficiency. McGarr (1984), has observed that seismic
efficiency decreases with depth.
There are a number of m1n1ng parameters that directly influence
the occurrence of rockbursts. Salomon (1983) has statistically
analysed these parameters, confirming the results for South
African gold mines.
il Excavation size; rockburst incidence increases with size of opening. ii) Abutment size; rockburst incidence increases with decreasing abutment size up to a critical width at which point the incidence
19
c
1 1
decreases. iii) Depth below -surface; r'ockburst incidence increases with
1 depth. 7-
iv) Dykes; rockburst incidence increases lin the vicinity -of dykes and faults. Mining through dykes and faults at acute angles also increases incidence.
3.2 MECBANICS OF DESTRESSING
The prevention of rockbu~ts by destressing wa~ first introduced by the South Africans i~ 1952, Roux et al. (1957) and Hill et al. (1957), for controlling rockbursts in the deep gold mines. Initî81 results, as discussed in their papers, werê very promising, however, ~s mining continued to greater depths, bursting
persisted. In rece~t years, with more sophisticated design and evaluation tools, destressing has proven to be a very effective r
method for rockburst ~revention. This has generatep much renewed interest in the method.
Figure 3-2 shows the initial theoretical basis behind destressing.
The concept came about from the observation of the natural fracture zone surrounding underground excavations~ Rockbtlrsts
" seldorn occurred in this fractured zone and it appeared to offer a
cushioning effect to any bursts oc~urring beyond it. It was
postulated that increasing the depth of this fractured zone rnight
reduce the occurrence of rockbursts and reduce their violence.
The fracture zone is increased by drilling hales ahead of the face, or into a pillar, and blasting them with explosives. The
hale spacing and explosives loading must be such that the ground
is fractured, but not broken to the extent that mining ia no
longer possible. If blast fracturing is conducted before mining
commences, the practice is termed preconditioning. If the blaat
fracturing is conducted once rniping has cornrnenced,- and the ~a ia
under high stresses, the term destressing is used. This thesis deals primarily with destressing. Both methods alter the rock
~
rnass rnechanical properties in similar ways, although,
preconditioning the rock prior ta it being critically stressed
20
............... --_ .. _.~----:::---------------:-----------------
Flour. 3-2: The InitiaI Concept Of De.tre •• lng
/
/ /
/
( DIstance from face
.. unfractured rock
After Roux et al. 1--857
c
21
:)
will result in somewhat different behavior. The d~cision to precondition is usually based on computer modeling or past
experience. The decision to destress is general~y made after
visual confirmation of high stress conditions and proneness to
bursting.
The mech~nics of destressing are not fully understood as yet.
There are a number of observed and postulated phenomena
contributing to the successful application of destressing. ,
i)Reduction in stiffness due to fracturing: \
a) causes redistribution of high stresses; to be effective
the,stresses must be reduced below the strength of the
structure, '
b) reduces stiffness of mine structure below that of mine
loading syst'~.
c) causes convergence.
ii) Energy dissipation by fractured rock:
a) energy is dissipated in closing new fractures, this
reduces energy accumulatton,
b) energy released beyond the fractured zone will be
absorbed by it.
iii) Transition from brittle elastic to pseudo-plastic:
a) fail~re of pseudo-plastic material is non violent.
iv) Reduction of peak and residual strength due to fracturing:
a) causes convergence.
c) causes failure and redistribution of high stresses.
v) Triggering a rockburst:
a) the destress blast may induce the release of energy by
initiating a rockburst.
To be effective, stresses must be transferred to solid rock where
they can be contained, stress transfer that destabilizes an
adjoining area has no benefit. Destressing to date, has been
utilized with success to prevent strain bursts and pillar bursts.
To this author's knowledge it has not been utilized on fault slip
type bursts.
22
(
(
3.2.1 BEBAVIOR OF FRACTORED ROCK
Cracks and pores have long been recognized to influence modulus
and poisson's ratio, Walsh (1965a, 1965b, 1965c). Increases in
modulus and Poisson's ratio with increasing stress, non linear
stress strain curves and hysteresis have aIl been explain~d by
the presence of pores and cracks.
Uslng the reciprocal theorem of Love (1927), Walsh derived
formulas for the effective modulus, poisson's Ratio and
compressibility of fractured rock. The formulas required
knowledge of average crack length and concentration, values that
are generally not available. The work concluded that modulus and
poisson's ratio are related inversely to crack concentration.
Compressibility is proportional to crack length to the third
power. Under high pressure, when aIl cracks are closed, the
effective modulus is equal to that of uncracked rock. The
effective poisson's ratio may exceed the uncracked poisson's , ratio at high pressure, this is the result of cracks opening
paraI leI to' the direction of maximum stress. Rocks with low
effective modulus exhibit high poisson's Ratio.
Lama (1974) demonstrates in model studies that deformation
modulus and uniaxial strength reduce with increasing nurnber of
joints. The modulus reduction becomes smal-l after the number of
joints exceeds six. For horizontal joints, the maximum strength
reduction was found to be 30 to 40 percent. For vertical joints
the strength reduction was 60 to 70 Bercent of the intact rock
strength. Studies using cubes to simulate orthogonal jointing
resulted in the following relationships.
E = k + ( Lii )B
where k = deformation of model containing more than
150 joints, the limit nurnber beyond which
no further reduction in modulus is observed
fi = constant
23
1
,. J"
- L = length of model
l = length of element.
~odulus value and uniaxial compressive strength were observed to ~
vary with the orientation of the joints to the loading direction
and joint continuity. Both modulus and strength values decrease
as continuity increased and angle between jointing and loading
direction approached 30 degrees.
Failure of the composite models was progressive. As the number of
individual elements failed, deformation increased. Post failure
curves were flatter for models containing a large number of
.elements and failure was less violent.
3.3 NOMERICAL MODELIN~FOR STRESS ANALYSIS
Numerical modeling to analyze ground stresses and displacernents , is a well recognized and accepted technique. Most modeling 15
now performed utilizing high speed computers. The aécuracy of
any modeling depends upon the sophistication of the computer code
and accuracy of the input parameters. With simple/codes and j
reasonably accurate input parameters (rock properties, geology
and in situ conditions) it is possible to obtain'results within
the correct order of magnitude. This type of model is easy to
use and quick. A more accurate prediction, as is required for
destressing, requires a more sophisticated code and very accurate
knQwledge of the input parameters. Model calibration via back
analysis is very important. Setting up and running these models
ta~es a greater length of time. This section briefly reviews
numerical modeling and its application to destress blasting.
withthe advent of
extremely popular.
of their high cost
computers, numerical methods have become
Physical models are seldom used now because
and time requirements. Likewise, photoelastic
studies are impractical. Multiple analysis of different mining
or destressing scenarios is now possible, in a very short time
24
\(
(
period, to d~termine the most appropriate course of action.
There are three types of numerical computer methods commonly
useg: differential, integral, and distinct element. A number of
advantages and disadvantages exist for each method.
With the difJerential method the entir~ problem domain is
discretized i~to elements. The outer boundary of the problem
domain is deff~ed arbitrarily allowing discretization errors to
occur throughout the medium. The differ~ntial method can be used
to model non-linear and anisotropie behavior. Rock masses
comprised of units with different mechanical properties can also
be modeled. Examples of differential methods are finite element
and displacement discontinuity. These types of programs are the
ones most commonly,! used 'for destress analysis.
With the ~ntegral-method only the surface of the excavation is ,
discretized; this greatly reduces the magnitude of the probl~m
and increases efficiency. The far field boundary copàitions are
modeled correctly and discretization errors are restricted to the
problem boundary. This ensures fully continuous variation of
stress and displacement throughout the medium. The integral
method is not suitable for analyses of non-linear or
heterogeneous behavior, as these negate the simplicity of the
integral ~ethods solution procedure. The boundary element method
is an example of the integral method. ,~ r
The third computational method is the distinct element method. In
areas of high discontinuity densities such as mine openings it
has been found that rock behaves as an assemblage of rigid blocks
that int~ract along deformable contacts. The model -area is
divided into a set a discrete elements corresponding to the
block pattern. The distinct element method assumes that block
geometry is'unchanged by the acting stresses, any deformation
OCcurs along the joïnt contacts which have definable mechanical
properties. Microscopie discontinuities exist throughout the
rockmass, however, they are disregarded because of their size in
o 2S
s
"
.. relation to the,area being mo~eled. Destressing introduces sorne
fracturing and enhances any existing near field discontinuities . . Although it may prove applicable, to the authors knowledge,
, simulation of destress blasting with the distinct element method
t has not been carried out.
Numerical models are useful for a nurnber of tasks related to
destress blasting. They are able to determine which areas are
under high stress and in need of destressing; they are able to
determine what reduction in modulus or strength is required, and
the aerial extent that should be affected; they use fuI for
evaluating the potential for creating new areas of instability as
a result of destressing; they are able ta predict _the most
desirable location, ie. hanging wall, foot wall or within the ore
plane. A computer model for analyses of destress blasting must , meet a number of requirements, the most fundamental being that it . accurately predict the rock mass response to mining and destress
blasting. The practice used in most mîning situations is to
first c~librate the model to known rock mass responses via back
analysis. Once the model is calibrated it can be utilized with
confidence to perform the above mentioned tasks.
An i erative approach is best suited to determine what minimum
reduc~'on in modulus of deformation or reduction in rock strength
is requi ed. Blast design to achieve the desired modulus i8
discu8sed in Chapter 8 and Chapter 9. Numerical modeling i8 also
able to predict the optimum destress bl,ast configuration. This . includes the required volume of ground to be blast fractured nnd
-location ie. hangingwall, footwall or within the ore plane.
\J Invariably there is stress transfer from the destressed zone ~nto
thesurrounding mine structure. Numerical models can be used to
;nalyze this transfer. If a new area of instability occurs,
appropriate action such as redesigning the dfstress blast can
take place.
26
(
(
Destress blasting reduces both the.modulus of deformation and
peak and residual strength. An~lysis at CRLM, Scoble et al.
(1987), indicated that with a finite element program, destressing ,... could be simulated as a reduction in modulus of deformation.
Hansen et al. (1987), reports that reduction in modulus does not
result in good simulation of destress blasting with the computer
program NFOLD, a displaèement discontinuity model. It was found
that convergence prediction with this model wa~ strongly
dependent upon peak aod residual rock strength and only
marginally dependent upon ~lastic modulus.
3 • 4 'l'HI EXPLOSIVE MBCBANI SM AND ROCK fRAC'l'ORING
..
An explosion is the result of the chemical reaction of a thermodynamically unstable material. The reaction i5 divided
into two zones: the reaction head, followed by the reaction zone. 1:1 '
A complete discussion of the chemistry and thermodynamics of the . reaction can be found in Cook (1970). The velocity at which the
reaction occurs, velocity of detonation (VOD) , is the most
important parameter governing the violence of the reaction. Two
.types of reactions are possible: explosive detonation or
explosive deflagration. Explosive detonation is characterized by
h~gh YOD, high explosive pressure, high detonation press~re, and
rapid release of heat. The detonation front travels at a
velocity greater than the seismic velocity of the explosive
material. The explosive reaction occurs in, and in turn supports
a shock front. Explosive deflagration occurs when the YOD is
- lower, resulting in lower pressures. Only detonation is of
interest in the process of fracturing rock.
As the explosive reaction proceeds a quantity of energy is
transferred to the rock. This is knQwn as the dynamic pressure
pulse and is derived from the detonation pressure. In close
proximity to the blast (rIa < 2) this pulse can be approximated
as single spike or step pulse, Figure 3-3. The pulse radiate out
in aIl directions but is,generally only considered in the
27
" • •
-.
o
Flgur. 3-3: Normallzed Dynamle Str ••••• : Step Pul •• tnd Wav. Train
.4
o
- ·8
: .. ~"""''' ;
... i .-.- \
.... \J ,.../ ...,
:
radial stress
1S
r:: dl.tanc. from bl.at h~'. • :: bl •• l ho'. r.dlua
28
time: ... ec
after Mohanty (1982)
-...,.
c
c
c
j
f'
tangential and radial directions. With ~me ~d dista~ce the wave separates into frequency components and d~fferent wave bype~. At a large distance,- ria > 10, the wave can.be approximated as a damped sinuso~dal wave train, see Figure 3-3.
Two processes are beli€ved to govern the transfer of the exp~osive pressure to the rock: stress wave transmission and borehole diration. Wave transmission from the detonating
explosive is governed by impedance matching of the explosive and
rock. The impedance relationship as weIl as an estimate of the -
explosion detonation pressure are given by:
where ~t = transmitted st±~ss Pi =--pressure in detonation front
d2C2 = density and -seismic veloci ty of rock -,
--d1C1 = density and VOD. of explosive
Using an explosive which provides the bes~ impedance match will
~heoretically result in the greatest stress transmission, although achieving even a close match is generally not possible~ due to the vastly different density and seismic~velocity-values.
At CRLM destress blasting is conducted with AN/FO in 45 mm holes,
rock density is 2.5 9 per cc. Assuming an explosive density of 0.9 g per cc, a seismic velocity of 6310 m per sec in rock, and a
VOD of 31QO m per sec, the resulting detonation pressure is 2.2
GPa. The transmitted stress would be 700 MPa.
The second and probably more important method.of stress transfer
is oorehole dilation. As the detonation front travels down the
borehole it causes hole expansion and contraction. Thie
29
'oscillat~on generates stress waves in the rock. Favreau (1969) _ has shown that the generation of strain waves in the rock is a function of ~ensi~y, Poisson's ratio, Young's modulus, blast hole radius and explosive properties.
At the periphery of the blasthole, energy transfer appears ~o be in the form of a shock wave. The pressures a'r~ not great-enoug~ ~ to create shock waves that travel at·velocities greater than the seismic v~locity, however, the pulses are of very,high energy content. The rock is incapable of transmitting this energy so crushing occurs,-Which causes rapid attenuation of the energy.
Once the wave energy has attenuated sUfficiently, an elastic or 1
seismic wave carries on out into the rockmass, Initially, the t."
strain induced by the elastic wave is still high enough to create fractures. Fracture growth velocity is ~proximately one third that of the seisrnic velocity (speed of the ela.stic wave) , as such fractures can only extend while they are within the stress pulse. This results in discontinuous fractures. The fracture velocity at 'CRLM is estimated to be between 1500 ahd 2100.m per se~ond.
Measurements of the wave energy away from the blasted hole indicate that only 10 "to 20 percent of the total explosive,energy
ois radiated as a seismic wave, the remaining'energy is used in crushing, fracturing, heating and noise. Duvall and Stephenson (1965) have suggested that sorne of this sei smic energy is· attributed to a change in potential energy due to enlarging the borehole. Their calculations indicate the amount to De-small unles~ \mder a very high stress regirne.
,The expldsion aiso ~reates a-Very high temperature and pressure gas. The gas exerts a quasi-static pressure on the surrounding rock. The pressure r~mains until the gas can escape through
, "..--
cracks or out the borehole collar. The gas' pr~ssure is much lower than the detonation pressure, this researchers'to question its significance Although it seems unlikely-that the gas
30
has lead Many -in the blasting process.
pressure is great--enough
c
c
c
to create new fractures, it is probable that e~isting fractUres may be extended or connected together. Obviously in situ stresses must be exceeded along with t~sile strength of the rock before additional damage can~ccur. Kutter and Fairhurst (1970) analyzed the explosive and gas pressure as two separate components. Their work indicated that gas pressure is important only if the cavity is first pre-conditioned by the exp~osive stress wave.
3.4.1~RACTORE CREATION
Fracturation occurs when the stress exceeds the strength,of the rock. \n the case of blasting~ when the amplitude of the ela-stic ,stress wave exceeds that which the rock can transmit. A single crack advancing at a constant rate is known as stable crack growth. If t~e- amount of strain induced by the stress wave
oQ
exceeds a certain limit, unstable crack growth occurs. Unstable crack growth is characterized by bran ching and multiple crack extensions ."
The stress wave can be resolved into radial and tangential . directions. Both directions have compressive and tensile components. On a theoretical basis it is found that the tensile com~nent o~ the tangential stress is very high, resulting in radia\ fractare$~ With time and distance the radial stress ,
develops a strongir tensile component resulting in tangential of hoop fractures. The compressive component is commonly assumed to have little influence on fracturation due to the high compressive 'strength of rock >'_
Theoretically the radial and tangential stresses can be calculated at any ,point in the rock mass. ' The simplest calculation is that of the classical stress analysis °of a thick wall~d elastic material subject to internaf.and external loading. The equation ~s given below, its theoretical basis' can be found
1
31
J' i , .' ,
~in most texts·on stress analysis such as Tirnoshenko and Goodier (1970) .
U rr = 2 (P~d2 - Pir2)
-p------------- + ------~-------------1
(d2 - r 2 ) a.? (d2 _1"r2 )
~ (p d 2 - p.r2 ) ee - -0 ~
where urr = radial stress uee = tangential stress
r = intern?l radius of cavity (blasthole radius) d = radial distance to outer boundary of cylinder
Pi = internaI pre-ssure generate by explosive • Po = extern~l pressure (in situ stress)
R = radial distance to point of interest
SolutJon of the above equations for a destress blast at CRLM
suggest that the maximum radial distance for compressive failure to occur is 0 .. 09 m (3 hC!>le diameters). Tensile failu,re should not· extend beyond 0.14 m (6 ho}e diarneters). The analysis -assumed a static pressure was applied and failure is through intact rock, not along the weaker~re-exi~ting discontlnulties. Input parameters were r = 22.5 mm, Pi = 1.6 GPa, tensile strength of rock = 15.6 MPa, compressive strength = 125 MPa.
, An,approximation of,fracturation can also bè determlned by stress and strain determined through peak particle velocity relationships. If the stress wav~ is assumed to be sinùsoldal and plane, the following equatlons can be derived for stress and strain. There derivation can be found in Dowding (1985).
~(f = d cp up up-:cp d Cs Us
32
(
/
(~
(
where \.
-
, q = compressive or tensile str~ss l ,-.. E = strain T = shear stress v = shear strain
cp = primary (compression) wave seismic ~elocity up = primary wave particle velocity Cs = secondary (shear) w~ve seismic velocity Us ='secondary wav~ particlê velocity
Negative strains are considered compressive. A plane compressive wave produces an equal magnitude s~ear strain, however, for equal particle veloc~ty, shear stress will be greater than longitudinal stress due to the lower seismic speed of t~ shea; wave. This method requires accurate prediction or measurement of peak
particle velocity close to the borehole. . . Prediction of fracturation is made more difficult when
consideration of in situ stresses is necessary~ as is the case .. with destress blast±ng. Most theoretical models do not take
1
conf·inement or in situ stresses into account, as their effect on 'Il' • •
the fracture process is not ëasily quantified. If the rock is considered to be under no confining stress, blast induced failure
will most likely be tensile. The tensile component of a blast
stress wave is less than the compressive component, however, the tensile strength of the rock is typically only 10 percent of the
compressive strength. If the rock has an in situ stress imposed on it, it may not be possible to fracture it in tension.
Assuming the principle of stress supérposition, the ténsile
- component pf the blast wave may not be sufficient to overcome the compressive force and tensile strength of the rock. When the
compressive stress of the blast is added to the compressive in situ stress, it may be sufficient to cause compressive failure of the rock. It is noted that compressive strength increases with increasing in situ stress and confinement. The extent of blast
damage""is depe~t upon the in situ stress conditions; it is
33
o
conceivable that under certain in situ stress conditions, blast
damage will not occur where it otherwise would be expected.
3.4.2 PREVIOUS STUDIES' OF FRACTURE EXTENSION
D~e to the complex nature of the detonation process and the
cesponse of rock to blasting it is extremely difficult to predict
fracture extension. Experimental and analytical'studies of crack
growth have been undertaken in the past with variable results.
To this· author's ~Owledge"no theoretical method exists that
satisfactorily p~dicts fracturation. Researchers routinely
, conduct small scale tests on laboratory samples, however, scafing results has proved to be very difficult. A number" of in situ:
investigations of blast induced fracturation in confined ground
have been undertaken. The reasons for the investigations and the
type of rock used vary considerably making it extr ely ifficult \
to compare studies. Fracturation is both site speciffic d
dependent upon the explosive and blasting arrangêmen~. The
method of fracture detection may also affect results_
The rock surrounding a blasthole can be divided into three zones,
Figure 3-4. A highly fractured or crushed zone surrounds the '.
hole. This typically extends to one hole diameter. The
shattered rock may be blown out of the hole by the escaping \
detonation gases. The transitional or non linear zone i5'
typified by a distinct drop in the peak stress, however,
extensive radial fractures still persist, Atchison and
Duvall (1963). The elastic or outer zone contains an ever
decreasing number of fractures. It theoretically extends
indefinitely, however, interest is placed only in the area
containing blast induced fractures. The extept of blast induce~
fractures in the linear zone is of primary importance in the .~
design of destress blasts.
Kutter and Fairhurst (1970). analytically determined that crack
extension should be six Qole diameters for a spherical charge and
34
, Figure 3-4: Zon •• Surroundlng Bla.t Hoi.
• /
\
\
c 1
\
"-
after Du Pont 1977
35
~'" '" . .,'
nine hole diameters for a cylindrical charge. Their-paper also
concluded that pre-existing fractures grow preferentially to new
ones and create a crack free zone in their immediate vicinity.
Fractures grow preferentially in the direction of maximu~
principal stress, however, in situ stress may suppress iracture
propagation in one direction in favor of crack initiation in
another. The extent of fracturation is dependent upon the , ,
strength of the rock, energy output of the explosive and energy
absorption of the rockmass. Increasing the charge beyond a
certain limit will result in a greater crushed zone but no
increase in the extent of fracturation. This conclusion is based , \
on th~\assumption that there is a limit to the amount of energy
that rock can transfer. If the limit is exceeded, crushing will
occur. Once enough energy has been attenuated, an elastic wave
will travel on outwards.
Olsen et al. (1973) studied fracture extension around confined
blastholes in a competent granite. The studies used a high
strength explosive. Microfractures were observed, in thin
seétions, to extend to ten borehole diameters. Siskind and
Fumant! (1974) observed fracture damage to seven borehole
diameters. Their work was carried out in granite using AN/FO in
165 mm holes. Siskind et al. (1973) studied fracturation around
blastholes in shale pillars of an underground mine. Using AN/FO,
fractures were observed ta extend nine hole diameters, using
dynamite, fractures extended ta twenty four hole diameters.
Steckly et al. (1975) studied confined blasting for the purpose
of in situ leaching. They observed fractures to ten hole
diameters about 230 mm diameter blast holes. Fractures
preferentially occurred along pre-existing discontinuities.
3.4.3 BLAST HOLE INTERACTION
Kutter and Fairhurst (1968) theoretically studied stress wave
reinforcement and interaction. The four possibilities that they
studied were as follows:
36
(
(
1) The delay betweèn blasts is so long that the pressure -from the first blast i8 negligible by the time the second charge is initiated. i
2) The delay between blasts is such that the quasi static pressure' from the first blast is still significant when the v?~ond charge is initiated.
3) The charge in the secopd hole initiates as the dynamic wave from the first hole passes over it.
4) Simultaneous· initiation.
The four cases are illustrated in Figure 3-5, where two blastholes
are considered.
When the delay between detonation of two charges is,long, no
interaction between stress fields occurs. If the blast holes are ,
sufficiently separated, the situation can be considered as the
explosion of two independent charges. Typical destress blasts
at CRLM use blastholes spaced at 27 to 33 hole diameters. If a
1.6 GPa stress wave is assumed to be generat~d, and attenuation
of the dynamic str~ss is proportional to r-1 . 5 , stress levels ,
on adjacent holes will not be great enough to cause any
modification of the blast hole periphery .
. If the charge in hole 'b' detonates when the quasi static
pressure exerted by hole 'a' is still significant sorne stress
interaction will occur. The pressure from hole 'a' will cause
stress concentrations about 'b'. High tensile stresses will form
at points i and ii, and high compressive stress at points iii and
iv~ Figure 3-5. ~uperposition of the stresses produced as hole
'b' detonates will cause preferential fracturation in the
direction of the plane of the blastholes. Fracturation of this
nature is desired in piliar dest~ess blasts. The fractures in
the plane of the blast holes may also be further extended by the
gas pressure from h9le 'b'.
When the dynamic explosive pressure from hole 'a' passes hole 'b'
as hole 'b' detonates, the results will be similar to case 2.
High tensile stresses are created at points i and ii.
37
\
Figure 3-5: Four Ca ••• , of Initiation Delay aetw.en Charg •• '.' end 'b'
hol. 'a' hol. 'b'
CASE 1
r 1
... ..1 "- .. r
-"lI-
a 1
1 OU."-lltllle Drn,m'e
* r
CIISf. :5
a
~O l QUOI'·"allc:
-,
r .. .; A
CIISE 4
,Outll·llolle
r -
...
after Kutt.r and hlrhur.t 1887
o
38
(
,
If the two holes detonate simultaneously, maximum stresses, due to superposition, will occur at mid point between the holes.
Unless the 'ho les are very close, it ~s not likely that the stress
levels will be great enough to create new fractures between the
holes, although, existing fractures may be extended.
~he previous discussion has assumed negligible in situ stress.
If the in situ stress is hydrostatic there is no preferential
stress concentration around hole boundaries. Preferred fracture
direction is unaffected although the extent of fracturation will
change.
When the principal in situ st~ess is perpenditular to the desired
fracture plane, problems are encountered. Tensile stresses are
created which favor the formation of fractures in a direction
parallel to the principal stress. It should be noted that the
actual hole periphery has no stresses acting on it as it is a
free surface; stress concentrations occur a short distance from
this boundary. The problem can be overcome by imposing a stress
field 90 degrees to the in situ stress. This stress field is
conveniently applied by the detonation pressures from an adjacent
hole. The best results would occur if detonation of hole 'b'
occurs as the dynamic pressure wave of hole 'a' is passing over
it. Analysis of the usual destressing practices at CRLM suggest
that the in situ stresses would not be overcome even in the most
Ideal situation. Assuming in situ stresses perpendicular to the'
desired fracture plan are 50 MPa, blasthole spacing is 33 hole
diameters and the stress wave attenuation is proportional to
r-1 . 5, the minimum transmitted detonation pressure required to
overcome the stresses around the adjacent borehole is 543 GPa.
Typical explosive pressures are weIl below this.
In the case of both in situ and no in situ stress it is most
desirable to have the explosives detonate as the dynamic pressure
from the adjoining hole passes over it. As discussed above, in
situ stresses may not be overcome, however, for destressing
purposes it is desirable to create as many fractures as possi%le ~
, 39
o
o
even if direction can not be controlled. Knowing the seismic
velocity of the rock a delay sequence can be determined to
optimize fracturation. Delay scatter, however, will result in r
inadequate results. Pressure sensitive caps would provide the
most desirable initiation, however, the inherent safety risks
make them unsuitable. Sinee cap timing scatter ls quite
signifieant, particularly in higher delays, the author believes
that blasting aIl holes the same delay will produce satisfactory
results with adequate safety.
A longer duration quasi-static gas pressure will more readily
results in constructive stress interaction. In order to enhance
the effects of the gas pressure on fracturation, it is
recommended that stemming be used to prolong the duration of the
pressure.
An important factor which has not been considered in this
analysis is the influence of pre-existing-discontinuities. The
stress level required to activate fractures along pre-existing
defects is mueh less than that required to initiate new ones. At
CRLM major discontinuities generally run parallel to the
direction of desired fracture propagation. Studies at the mine
have shown that with the exception of the crushed zone around a
blasthole, blast induced fracturation forms along pre-existing
discontinuities, Scoble et al. (1987).
Stress interaction occurs in three dimensions, not just in two as
discussed above. Stress waves originate along the entire length
of the blast hole and radiate out jn aIl directions.Fracture
development is alsO a th~ee dimensional phenomena, not two
dimensional as is commonly assumed for modeling purposes.
Mohanty (1~82) has shown that based on dynamic tensile failure
criterion, 3 orthogonal sets of cracks should occur. Two
orthogonal radial sets, one horizontal, one vertical, arise ~rom
the tangential stresses. The third set is circumferential and iB
the result of the tensile portion of the radial stress component . .. ,
'The circumferential cracks do not occur until the tensile
40
•
j
Plgur. 3-8: Thr •• Dlm.n.Jona' Fracture Development. ,
BOREHOLE
( .fter Mohanty 1982
i
, ;
"
c
\
o
cornponent of the radial_stress is sufficiently large. Figure 3-5 illustrates this predicted fracture scherne. As discussed above, in situ stress and pre""existing discontinuities will signif~antly alter this idealized pattern .
. 3.5 NOMERlCAL MODELING FOR l'RACTORlI: ANALYSIS -,
MOQeling of blast induced fracture extension is still in the
development stages. The response ot rock to blasting is very - difficult to predict due to the êomplex_interaction between rock
and explosives, as well as the complex behavior of the rock and _ e~plosive individually. As with aIl numerical rnodeling, the
accuracy of the results ls dependent upon the sophistication of the model and.accuracy of the input parameters (explosive properties, rock properties, blast hole configuration, etc.).
Most avai~able codes for nurnerical modeling of fracture extension
use established elastic theory and static loading conditions.
These codes make a number of assurnptions such as linear elastic
behavior and a homogeneous, isotropie matérial. Usually, only
tensile failure is considered. Because of greater explosive use,
and less complicated environments, most blasting simulators are
designed for use in surface operations. As such, the influence
of in situ stress is not considered. This severely lirnit~ the
confidence placed on results obtained from these models for
ponfined blasts. It is hoped that more sophisticated codes,
suitable for modeling confined blasts in high stress regimes will
bec orne available in the near future.
(;,)
42
c
4 REVIEW OF DESTRBSS BLASTING PRACTICES
The prevention of rockbursts by destress blasting was conceived in South Africa in the 1950'5. The concept came from the observation of the natural fracture zone surrounding underground
ex~avations. Rockbursts seldom occurred in this fractured zone which also appeared to form a protective cushion against bursts
occurring beyond it. It was postulated that increasing the depth
of th!.s fractured zone Iflight reduce rockburst frequency and
intensitYi the high stress region wouid be shifted farther away . from the stope opening and the extent of the protective cushion
zone would be increased.
~
FOllowing the South African initiative, man y m1n1ng camps with
rockburst problems initiated destress blast programs. It was
quickly realized that the technique could aiso be used in
development headings such as drifts, raises and shafts, and aiso
for destressing large areas of groupd such as sill pillars and
entire stopes. The techniques used have varied greatly, depending
on stope geometry and engineering approach. Although much
destress blasting has been perforrned, documentation on the
practice is sparse. An unpublished study by the Ontario Ministry
of Labour, Frantzos (1980), reported that 18 Ontario mines had
used or were using destress blasting for ground control. Most of
the available literature reports on successful destress blasting,
however, not aIl destress blasts have achieved the desired
results, Quesnel (1987), Oliver et al. (1987).
'During the course of study for this thesis, site visits were made
by the author to Kerr Addison and Macassa mines where destress
blasting has been used extensively. These mines have geological
environments similar to Campbell Red Lake Mine (CRLM). This
chapter presents detailed case studies of destress blasting at
CRLM, Macassa and Kerr Addison mi~s. Indicators llsed in the past
to determine when to destress as weIl as sorne of the reported
benefits of destressing are discussed. A review of case studies
found in the literature is also included.
43 .
o
4.1 INDlCATORS l'OR DBTERMINING WHBN AND WHBRB TO DB1JTUSS
Many--different techniques have been~used at mines to determine when ~estressing is ~ecessary. The decision-to des~ress is made once conditions indicate the onset or expected onset of higb stresses.
1 -./
Present day numerical modeling methods can accurately predict 'u
stress conditions, which in turn can be used to determi~e when and where to destress. In the past, stress prediction was not readily determined. A number of indicators and criteria used previously to determine when and where to destress are listed below along with referénces to the available literature. P~st
experience was often used as a guide; engineers would decide to' destress knowing that similar conditions had resulted in rockbursts on c:ither occasions. The occurrence of seisffiic' activity and small rockbursts also indicated the need to destress an area. Rock appearance has been used to identify ,over-stressect rock, however, the method is not weIl defined or easily quantified. Deform~tion of openings such as drill holes incticates high stressés; an estimate of magnitude and direction can ,. also he determined. In most cases more than one indicator is reviewed prior to making the decision to~destress.
Preyious experience; Hill et al. (lES7), Jianyun and Jiayou . (1987), Karwoski et al. (1979), Cook and Bruce (1987),
Makuch et al. (1987), Blake (1979).'
~ occurrence Qf seismic actiyity, ~ rockbursting:
Quesnel and Hong (1986) 1 Morrison '(1964), Moruzi aI)d Pasieka (1964), Oliver et al. (1987), Dickhout (1963),le Bel et~al. (1987), Slade (1968).
~ appearancej Issaccson (1962) reports a pillar hav~n9 a monolithic appearance, Roux et al. (1957) reports overstressed xock having a glassy appearance.
44
t
L ~ \
(
, , , ,
l
\
c
r
Tunnel deformation ~ drill ~ closurej Lama (1972).
~ mgàeJlng: Makuch et al. (1987), Karwoski et al.
(1979), Blake (1972):
..
, 4'.2 BZNZI'ITS AND RESULTS OF DIlSTRBSS BLASTING i,
• Successrul destress blasting causes a number of eff.ects on the rockmass. The following list of ,reported beneÏits is derived'
l' from the references sited in the above"'section.
- 'reduction in the frequency and magnitude of seismic activity 'i
, and rockbursts in the destressed zone. J ~h~ occurrence of seismic act:ivity and rockbursts in adjacent lt
, 1 1
. blocks of solid ground. -,closure.
- reduced deformation. reduction in stress.
- overall improvements in ground conditions. -_reduced scaling.
l ' - :reduced overbreak. ~
,1 >1
,- ,\reduced rockbolt tension. -- ~ncrea'sed fract~ring-. '
J 1
:
In sorne cases c'ontradicting results have beenQ reported. Lama (1972), reports reduced deformation in a tunnel after fi ring radial destress holes; most authors report an increase in closure after destressirig. This-discrepancy i~~elY due to differences in application and geol6gidalr~nvironment, however, an accurate .-. interpretation ls not possible because of limited information. Both reduced overbreak and scaling as weIl as increased fracturing have been reported in the s~me paper,' Dickhout (1963).
This author provides no explanation for 'this apparen~ , '
contradiction. Fracturing has been reported to increase by as _. muc~ aS 100 percent, Dickhout (1963), Moruzi and Pasieka (1964)
report a more modest 15 percent increase in fractures. Destress
45
· holes are generally observed to have enlarged after blasting. Collars are often bell shaped. Harling '(1'962), reports a 75
'" percent increase in diameter along the length of the hole. t
4.3 CASE HISTORIES 4.3.1 SOUTH AFRICAN GOLO MINES
,
The first destress trials were conducted at 'the East Rand . Proprietary Mines in South Africa. The mine is typical of Soùth African .gold mines in that the gently dipping reef deposit was being mined by a longwall method.
œhe procedure ·used was to drill 51 mm diameter holes on 1.5 m centers, 3 m deep. The bottom half of the 'hole was charged with 60 percent strength Ammon Gelignite or Dynagel. The remainder ~f the hole was sternmed with sand. Holes were blasted routinely to
ensure a fractured zone extended at least 1 to 1.5 m ahead of the 'stope face. Holes were collared 0.3 m above the footwall and angled down 8 degrees# resulting in the toe penetrating 0.6 m into the footwall: In dyke material the hole spacing was reduced to 1 m and charging increased to 70 percent of the hole length.
-- . Table 4-1 summarizes the initial resu1ts. As discussed by Roux et
al. (1956) and Hill et al. (1957), the initial results were very promising. As mining continued to greater depths, however, bursting pérsisted and the destress program was abandoned.
Blake (1982') concludect that the early South African destressing did not provide a large enough fracture zone for the long stope faces and high stress levels at which mining was being conducted.
\.
Cook et al. (1964) report on the_ results of seismic monitoring of o
conventional and destressed production blasts in South Africa .. It was believed that in order for destressing to be successful
there must be a reduction in stored strain energy which would result in the creation of seismic energy. Monitored blasts' indicated no difference in the seismic energy"released by the two blast types; all energy was attributed to the explosives alone.
46
(
(
\
_ .. --,
T-able 4-1: Summary of Results of Destress Trials
Befor.e Destress~ng
No. of stopes_
Fathorns stoped 16,600 (j
Incidence no. of bursts 44 bursts per 1000
fathorns ,stoped 2.64
Sever!ty no. of severe bursts 22 %. severe bursts to
total bursts --.50 incidence of severe
bursts 1.32
Time '-bursts during day
shift % during day shift severe bursts day'
shift % severe day shift
Casualties killed
. injured total
Il / 25
5 110
5 24 29
/
After Destressing
17
17,300
29
1. 68
6
20.6
0.35
o o
o 1 1
Percent Improvernent
34
36
73
73
81
100
100 9"6 97
after Hill et al. (1957) .
..
t
47
\ '. .
J l
~
-4.3.2 COEUR~'ALENE NcrNIN~ DISTRICT
The Galena Mine, as weIl as many other mines in the Coeur d'Alene
mining district of Idaho, have successfully used pillar
destressing for a number of years, Blake (1982). Steeply
dipping, narrow, veins are being mined by overhand cul and fill
methods. Once ~he crown pillar is reduced below a 15 m thickness \ -
the ground becomes burst-prone. The procedure has been to
destress using 38 to 51 mm holes drilled from within the stope on
1.5 to 3 m centers, as shown in Figure 4-1. Blasting is with
high velocity explosives and milli-second delays. The blasts _
have resulted in substantial stope closure, reduced seismic
velocity and reduction in stress around the stopes. The degree
~f fracturing has not been such as to cause subsequent mining
problems.
Destressing a large block of ground, such as a pillar, has been
~--- ~xtended to entire stoping areas. Karwoski et al. (1979) report
on what they calI ground preconditioning at the Hecla mine in the
Coeur d'Alene district of Idaho. Large ,blocks of ground were
preconditioned or blast frac ured during level development, prior
to commencement of pro uction mining or the creation of highly
stressed, rockburst prone pillars. The object was to fracture
the ground sufficiently to prevent the creation of ~ighf.Y
stressed ground right from the start. Mining would begin and end
in fractured rock. Mining between leve1s, in areas outside the
preconditioned.zones still required conventional destressing. The
destress holes were fan drilled from crosscuts with maximum toe
sepa~ation of 2.3 m, Figure 4-2. Hole diameter was 92 to 102 mm,
the prima~y explosive was a high velocity water gel, charging as
indicated in Figure 4-2, aIl collars were stemmed with sand and
clay. Holes were toe and collar primed and traced with
detonating cord. Milli-second delays were used to reduce the risk
of air blast damage.
Preconditioning trials were conducted in both the vein and wall
rock. In both cases destressing was considered to have been
48
C ------.-
C'
c
\ l'
1 . ,
'1 '--------... J'/OO t...1 f .... ,.. .... r
1 i T r T 1'1 T 1 -T
"" .. 1 1 l ,1 1 1 1 1 1 1 1 1 1 1 1 1 1 1
0
.--1 • .. .. JO
! l g -. ....
!I S ltrnCIII A-A
Figure 4-1: Stope pmar Oestress Blast Pattern, Coeur O'Alene (
è
{ after" Blake 1972
Pounds of explosives
Tola/= 337
225 ~14
crosscut
o 10 20 1 l ,
.. ~. Seole ft Figur. 4-2: Pr.condltlonlng Bla.t Pattern Coeur D'Alene 1
after Karwoakl et al. 1979
49
l
~ f r -
effec~ive in reducing the anticipated high stress and burst prone
oondition. Based on drilling. costs, mineability'and results of
computer simulations, the on-vein preconditioning was considered
preferable to off vein preconditioning. u~
The preconditioning was monitored and evaluated with stress
gauges, convergence stations, seismic velocity surveys and a
microseismic monitoring system. A 28 percent reduction in the
seismic velocity and a red\ction in the stress level were
recorded in the blasted arra; this indicated that the ground was
successfully fractured by the blast. Stress and convergence
measurements during the course of mining agreed well with those
of linear elastic finite element computer modeling. The
microseismic monitoring continued to register activity in the
vicinity of the preconditioned ytopes as they'were mined,
although frequency and intensity were less than usual. Only
minor activity'was recorded within the preconditioned or shielded
ground. Mining between the destressed zones resulted in
increas~d seismic activity and the reoccurrence of rockbursts.
This confirmed the need to precondition the entire stope block.
As a result of the success of the first preconditioning trial, a
second, larger, preconditioning blast was cortducted at Hecla
Mine, Blake (1980). The block destressed was 137 rn long and 27 m
on dip. The drilling pattern and explosive loading was siroil~ to
the first trial. Drilling was exclusively in the more easily ~
drilled vein rock. Holes were 82 ,mm diameter, loaded to a 6 m
collar with Tovex 5000 cartfidge.' The holes were traced with
detonating cord. In order to minimize stress transfer rate and
air blast, the holes were blasted in groups of 5 or 6 at a time.
Raising up through the preconditioned rock resulted in only minor
microseismic activity. A second phase of this destress project
was planned to destress pillars on the over lying level. As
drilling was being conducted a major rockburst occurred,
destressing the area in question.
50
c
c
4 • 3 • 3 SODBORI APJ!.A MINES:
At the Falconbridge mine in Sudbur~, destressing has successfully t'
been used t~ alleviate rockburst prone conditions in steeply
dipping cut and fill stopes, Slade (1968). After a number of
rockbursts occurred in the stope hangingwall it was decided to use destress blasting. Initially 3 m long d8stress holes were
J
dri1led from within the stope out into the hangingwall and blasted with the advancing face.( As the pillar thickness was
reduced, 21 m long destress holes were drilled from the upper
level into the hangingwall ~s shown in Figure 4-3. The holes
were spaced at 6 m centers. Induced fracturing and loose were
observed i~the drift and stope. The stope closure rate
decreased after destressing and no problems were encountered when mining the pillar. 4
~oruzi and Pasieka (1964) report on a similar destress blast at
the Falconbridge Mine. The blast was monitored to\evaluate the
effectiveness of destressing. Blast holes were aga'in drilled
from the upper level into the hangingwall. Holes were 48 mm
diameter at 9 m centers. Drill core fractures, closure and
photoelastic shear strain were analyzed pefore and after the
blast. The closure rate was not affected by the destress blast. "
·There was a lS percent increase in the number of fractures
observed. The strain rate was obser~ed te increase slightly,
although only marginally above the accuracy of the measuring
technique.
f7 The Creighton mine in Sudbury has been destressing development
headings since the 1960's, Dickhout (1963). After the evaluation
of a number of blast designs the accepted practice became to
drill two 50 m,57 mm diameter, holes in the face as shown in
Figure 4-4. A 2 m overlap of subsequent destress holes was
always ensured. Blasting was performed with AN/FO. A 3 m collar
was left unloaded. Destressing caused a mark~à reduction in
overbreak as weIl as reducing the frequency of microseismic
activity. A 100 percent increase in the number of fractures
51
T."tSA" OfiSTR(S$ t!.QU
Figure 4-3: Stope Pillar Oestress Blast Falconbrldge
" PREVIOUS
OEST~ESS HOlE " OESTRESS HOlES " 6' 1:
-~~'!~-----~-----------_:~---------2S'± •
after Siade 1968
DRIFT
Figure 4-4: Development Headlng Destresslng At INCO
LOADED
I...-~----I
0AIWt1C \MTS / PLAN VlEW fOR HEXT ROUI4D \
+8 f LONGITUDINAl VlEW
52
efter Dlckhout 1963
Figure 4-5 Oév'elopement Headlng Oeatresslng At INCO
aUer Oliver et al. 1887 ~
(
c
c
visible in the drifts was reported. In one case a rockburst
occurred at the drift face when the overlap of destress holes was reduced to zero, Oliver et al. (1987). Destressing techniques
used at the mine have changed over the years. Drifts are now driven with jumbos: four destress holes are used as shown in Figure .4-5. The holes are drilled to twice the round depth, two
of the holes are angled up and out into the wall rock. Review of
the destressing practice revealed that the need to destress
depended on drift orientation and proximity to the orebody.
4.3.4 KERR ADDISON MINE
Kerr Addison Mine is located 40 km east of Kirkland Lake Ontario.
The mine has a long history of rockbursting and has used destres
sing extensively in the pa st to control the problem. Regrettably
very little documentation exists on the practices used at the
mine. The following section is based on a mine site visit and
personal correspondence with Mr. M. Plaunt, the chief mine
engineer.
In 1962 thé loss of the mine was feared due to rockbursting. Dr.
R. Morrison proposed a mining sequence and destressing program to
reduce the frequency and magnitude of the rockburst incidence,
Morisson, (1962). The. new sequence resulted in a 50 percent
reduction in production rate although the number of rockbursts
also necreased. Destressing was never scientifically qualified
~lthouqh it was believed to have been an important tool for
ground control.
Initial destressing at the mine was performe~-in overhand cut and
fill stop~s as indicated in Figure 4-6. There is little record of
the effectiveness of this method, it is likely that any benefits·
gained were outweighed by the creation of a fractured back under'
which miners had to work.
53
•
e ... CIl
dealreaa holea
1
1 t -
Figur e 4-6: 0 es tres sPa tt ern St.ope Ba oks, Kerr A d dl s on MI" e
atler Frantzoa 1980
1
54
(~
c
Due to prodùction demands, a 46 m thick by 305 m long pillar was
created below 3100 level. The pillar contained long fIat backs
t in man y areas and was th~ location of extensive rockbursting.
Following Morrison's advlce a slot was mined through the pillar
with the aid of destressing. Destress holes up to 37 m long were
drilled from the stope and from 3100 level, Figure 4-7. The
holes generally followed the d~p of the ore although sorne were
drilled into the hangingwall. No explanation exists for the the
hangingwall destress holes except that they were reported to be
effective. This technique for bringing raises up to a level was
used ext~nsively throughout the mine after its initial success.
After the slot was mined through, square set stoping commenced in
an inverted "V" sequence moving away from the slot. Stopes ran
the full width of the ore zone but were restricted.to 15 m in
length. Destressing was performed routinely every second lift,
the pattern used is shown in Figure 4-8. Destress holes were
drilled with bar and arm type drills, 38 to 57 mm diameter. The
remainder of the ~re body was also mined with an inverted "V"
pattern between levels and an overall regular "V" pattern
throughout the mine. Destressing continued on a regular basis
every second lift. Rockbursts still occurred, however they were
less frequent and r~stricted to locations remote from working
areas.
No engineering studies were conducted into the effectiveness of
destressing. The opinion at the mine is that the method was •
successful and would be used again if a similar situation arosè.
The ground being destressed was blocky, and under high stress,
requiring the use of square set stoping. Any additional
fracturing caused by the destress blasting was not a concern. It ~)
is apparent that the blocks were tightly interlocked and only
loosened at excavation boundaries. Destress blasting would have
caused block loosening to a greater depth. The mining methods,
sequencing and destressing used at Kerr Addison were extremely
labour intensive and costly, however, they did prove effective in
combating the rockburst problem.
55
o
o
flll
o
hanglngwall ore zone
de .. tresa holea 1 1
.1 1·
3.100 level
tootwall
3250 level
. , ' Figure 4-7: Destre88 Pattern Ratslng Up To Level. Kerr Ad/dlson Mine
56
Figure 4-8: Destress Pattern Square Set St oplng, Kerr" Addison Mine
-
lead l .; -stope q) -
1 1
- ~-1 "1
>- 1 1 -"
1 < -
1 1" 1
,
.... 1 1 1
1
1 1
\ 1 1 1
-\
-. , 0;1 -~
SECTION V(EW
c l'
l -~ / hanglngwal
. 1 1
1 1 > 1 1
1 1
1; ~ '"
1 "
~~~
>"
l
1 1 - 1
\
L - -" ~
-"-l
footwall
PLAN VIE""
57
1
--
o
..... , ..,
1
!
4.3.5 KIRKLAND LAKE MINING CAMP
, Maeassa and Lake Shore mines have used destressing as a means of alleviating high stress conditions sinee the 1950's. the' following section is based on work by Harling (1965), Cook and Bruce (1983), Quesnel and Hong (1986) and te Bel et al. (1987), as weIl as a sit~Dvisit. The site visit was made to Macassa mine-and included-discussions with R. Hon~! ground control eng~neer for Macassa mine and W. Quesnel, chief ground control engineer for Lac Minerals.
Both mines are located ~n-Kirkland Lake Ontario. They,are Gwned by Lac MineraIs Ltd. At present only Macassa is operating. The entire Kirkland La~e mining,camp has a long history of severe rockbursting associated with areas of geological weakness, stiff rocks and high extraction ratios.
Longhole and short hole Q~stressing were practiced -at the mines in the pasto Harling (1965) reports short hole destressing to ~ave been eftéctive in reducing the bursting incidence"in breast faces and backs put gives no details. Quesnel (1987) reports that successful results are currently being achieved with short hole (10 rn) destressing used during driving of conventional
, --raises. Suecess is g~eatest ~hen holes are drilled in the plane of the rai se rather than angled outwards. Destress héle location~ is planned aecording to geological structure.
Longhole destressing was used as a st ope approached a mined out are~_or fault, in pillar recovery and to bring raises up to the leve:l-. Examples of actual destress blast hole ayouts are shown in Figures 4-9 to 4-12. In all these exampl he mine waa of the opinion that the blâsting was successfu shifting-high stresses away from the area. This was such_~yidenc~ as
reduced burst incidence and intensity, and the occurrence of . rockbursts in areas adjacent ta but not witnin the destressed ground. A more in-depth account of this work ia presented by Harling (1965).
f.
58
(
(--
fil "
fil 1
F.lgure 4-9:, Destress Panern For Plllar Recovery
\ destre88 holes
1
v-
\ \
after Harllng
\ Flgur,e 4-10:0estre8s Pattèrn For"Stop..a Approachlng Fault .,
~9 aUer Harllng
--
u
1
)
o deatress holes
fil 1
f "f.
Flgur~ 4-11: Destress Pattern For Approachlng Stopes
after Harllng
"" -=
Figure 4-12: Destress Pattern For Ralaes , ,
after Harllng
-
o . ~
,
c
Oéstress hole~ were usually diamond drilled 38 to 57 mm diameter. Percussive holes were used on occasion although deviation and
general_ho~~ quality were not as good. Holes were loaded with
AN/FO and collar primed. A 3 m collar was found desirable to
reduce cratering. In broken ground the hole -was traced with
Primacord to ensure complete detonation. Hole spacing was 7 to - 8.5 rn centers. This was believed to be.the maximum hale to hole
distance for successful dest-ressing. It has been reported, , , Harling (1965)'-, that a sign of successful destressing was 1;0
observe deformation due to rock movement in th~blasted hole.
The authors experience is that deformation and enlargernent is
~~mmon in most hol~s' afte'r blasting, particularly ~here many geological discontinuities and high in situ stresses existe Hole
deforrnation is not considered a suitable indication of successful '
7 destress blasting. 'When possible, the holes were drilled in the
vein: where not possible, duê to badly broken ground, holes were
·~led at 10 to 15 degrees into the hangingwall. Hole collars
were lined up êo as to minirnize damage to timber support. When /' '\
blasting, water Blowers were used to reduce the fire hazard and
bulkheads were used to prevent blockage of any travel ways.
Holes were blasted simultaneously. Ground conditions gen~rally
deteriorated, necessitating'further de~tressing, after advancing
two thirds of khe distance covered by the destress zone. In sorne
instances the mining of remnants had to be discounted dué to their low tonnage and grade
destressing. No rockbursts
'by longhole destres~ing ..
and the relatively high cost of
were reported to have been triggered
Although reports suggest the destressing was successful, this
author's e~perience suggests that the blast designs used in the
Kirkland Lake Camp in the past would only cause local fracturing
and modulus reduction in the,vicinity of the blastholes. Hole
spacing was too great to effect the rockmass between the holes.
61
\
1
1 Ma~assa mine is now working to depths of over 2000 m.
--Mineralization has become less scattered than in the past,
t'
resulting in higher area extraction ratios. Tne greater --. ~
extraction and greater depth is resulting in more frequent and
higher intensity rockbursts at the mine.
, .. In June 1986 a destress blast' was initiated at the mine in the
,58-40 stope crown pillar. The pillar is shown in long section in
Figure 4-13, a 55 percent extrac'tion rati.o had been reached. The
\ de<?ision to destress was based on a high incidence of seismic . '
, \ activity in the stope back as ~ell as indication of high pressure
by squeeze blocks. Numerical modeling later confirmed the high '"
stresses.
The vein material in this stope is a silicified basic syenite and }
tuff. The fqot wall is a less silicified basic syenite and tuff.
vThe hangingwall is basic syenit~ and syenite porphry. The vein
rock is the stiffest material. A major fault forms the
hangingwall in the w~stern section and swings over to the
footwall in the eastern~section. The fauit is tightly closed
with up to 2 cm chlorite gouge infilling. A sec~nd fault is
located 1 to 2 m in the 'hangingwall" parallcl to the vein . •
Diamond drill holes revealed a 3 to 4 m thic~ fracture zone
around the perirnater of the pillar. High stresses in the center ,
of the pillar were indicated by core discing.
The destress hole 'pattern is shown in Figpre 4-14. The blast was
designed with a powder factor of 0.045 kg/t. Due to hole
squeezing the final loaded powder factor was only 0.029 kg/t.
,Hole number 1, shown on Figure 7, could not be loaded due to a
break through. Hole number 14 was eut off and did not fire. To ( 1
, avoid vibration damage to nearby mine structures, blasting was
l~mited to 42 kg per delay. Collars were 3.5 m, stemmed with
C.I.L Inert Gel ànd barite. A high degree of eratering still
oeeurred around the blasted holes, this was due to the fraetured
-p~rimeter zone around the stope. Convergence points and à microseismie monitoring system were used to monitor the blast.
62
(
ri
, -' -
#3 shaft
58-40
r----' L-...-JI
J _ JO
5725 level
Figure 4-13: Long Section Macassa Mine
efte.,. Quesnel and Hong 1 sa6
63
.,
..,.
~
Ch ~
o •
":', ! ••••••• • f ••••••••••
:·~~mitmm . :~1~~n1HHH~
l~l~·i{tl: :~.::II.~.i·.'I·: \}f;l;(rl~i~iiji)l:~ \ll~,\;,r •. , ••• " ...... ' ...... t .... ' .••• ;>~ •• , ....... 1 ...... .: :.::,~.~.~.:.::. :.: .. ;.:: .• : t·:-: :." "." '/ " ~.?:::'.'~ ~'. 3 .:::::.:.:.:
, 1 1
\ 1 \
" \ \ \ 1 \
\
\ 58-40 Stope\
\
\ \
1 1 :::.:~
:.:·:·:V:·~lI~"\\".'~\.\.'.":':.:':':':',:':':':\';': ,
14
.f!
~j~~~~~tî~*~~~t~~~~:~}~~lf;~J~i ."::;:::~~~~:~~~,*~f:::~~:~~~:~ ~~;~:: ...... .
~1g~.~W~\j:: . ~t:!i~r{i;;l
13 12 11
L,~J .~. '\~
~i~;i:\~'~';~f11' ! fj'i:i'i,'~~~;;'~~:~f.~~;!!i~ii!·, t\ltll\l:ii:l~~;:,I·ll!~ft.~;i~~;I r:::::i}:;··:::·::.t:::!j::·:~:!~\!;::::i!::. 10 9 8 7 6 5 4 3 2
~
~ ... " ~' ... '.' ' ..
"
-/-J
Figure 4-14: Destress M,cassa
1 Stope
1 Crown Pillar Mine
after Q'uesnel and Hong 1986
c
(
Low and intermediate energy microseismic events occurred for
several hours after the destress blast. No events occurred within the solid block below the stope, suggesting it was capable
of sustaining any transferred stresses.
Convergence was recorded at aIl stations, the maximum was 28 mm.
The minimum convergence was 5.8 mm occurring adjacent to hole 1.
After a subsequent production blast a small rockburst occurred in
this area, displacing 25 'tons of rock. The convergence jumped to
33 mm. It is believed that the ground in this area was not
adequately destressed as a result of being unable to fire number
1 hole.
Drill core recovered from between destress holes 12 ~nd 13 was
not disced, indicating a stress reduction. Fracturing was
greatest at the location between the destress holes indicating
the creation of a continuous fracture zone between holes.
, Although only a ~ow powder factor was used, the ~nergy input was
sufficient to induce failure and interconnecting fractures in t~e
plane of the destress holes, this indicates that the stress on
the pillar was approaching the rock peak strength.
To date, no more major seismic activity has occu~ed in or around
the stope. The destress blast appears to have be~n successful in
transferring stress ta adjacent ground. The fact tiat one burst
occut-red in the vicinity of low convergence suggests that
convergence is a necessary effect of destressing.
!
During sinking of the No. 3(shaft at Macassa mine, rockbursting
becarne a continuous probleml after 1120 m; a number of remedial
measures were taken, including des-tressing. The following expert
from Le Bel et al. (1987) sumrn~rizés the methods and experience.
nA variety of blastholes drilling patterns was used to achieve destressi~g during shaft sinking. These included holes drilled into the sidewalls and loaded to their full length, inclined holes drilled downward into the corners of the excavation, and perimeter holes loaded with Xactex. Both sideholes aud corner holes tended to cause excessive blast-
65
--
induced fracturing and 100sening of the rock, subsequent rockbursts. successful in reducing
f
overbreak and, because of genera1 increased the amount of damage caused by Xactex loaded perimeter holes were not
bursts.
Below 5000 feet (1524 ml,' a systematic pattern of destress blasting was used. At the shaft stations (every 150 feet, 45.7 ml, four 2 inch (50 mm) diameter holes were drilled, using a longhole machine; for 90 feet vertically from the corners of the excavation (holes l, 3, 4 and 6 in Figure 4-15). In locations where bursting had been experienced above the station, an additional two holes (no's 2 and 5) were drilled. These holes were blasted with 3 taped lengths of Primaflex and were initiated with the first bench blast after the station.
Once excavation had passed the 90 ft. (27.4 m) longhole destress holes, a series of short, 12 ft. long, small diameter (1 3/8 inch, 35 mm) destress holes were drilled with each bench. The holes were located in the ben ch as weIl the end walls. Four short holes would be drilled per bench, i.e. holes 1, 2, 3 and 7 from o~e bench to the next.
H\tes \ and 8 had a collar of 5 ft. (1.5 m) of commercial stemming and ehe remaining 7 ft. (2.1 m) was loaded with Powermex. Of the various destress blasting methods used, the longhole destressing coupled with short holes proved to be least damaging,/, although it ls difficult to demonstrate any substantial
.. reduction in rockburst frequency." . ~
, 4.3 . 6 CAMPBELL RED LAKE MINE "vi'
-l. Destressing at CRLM'has included crown pillars in cut and fill
stopes, remnant boxhole pillars in shrinkage stopes, and face
destressing in horizontal cut and fill stopes. In general ,
destress blasting is deemed to have been successful. In order to
gain a better understanding bf the mechanics of destressing the
most recent crown pillar destress blast was heavily instrumented
and monitored. This section is based on mine site visits during
destress blasts and discussions with the following: T. Makuch,
N. Neumann and B. Lokhorst - mine engineers, Dr. W. Blake - rock
mechanics consultant, Dr. D. Hedley - research Scientist at
CANMET. Information on the geology, layout and rockburst history
of the mine can be found in Chapter 3.
66
\
~
c
c
c
~ .;
)
r
..
12 ft
• 1
4 •
- -.... 8
• 2
5 •
7
• 3
7 •
~--,
~;;. ......
-\ .,
Figure 4-15: Shaft Slnklng Destress Pattern, Macassa Mine
•
-7
after le Bel et al 1987
67
..
.,
., , r ..
\
The first cut and fill pillar destressed was in 1102 E. 'B'
séope, located in the 'A' ore zone on the 10 level. The area ~ad a very high extraction ratio and was subject to considerable
bursting and spalling. Destress holes were drilled in a fashion
similar to that used in the Coeur d'Alene district of Idaho. ,Holes were 45 mm diameter, at 1.8 m centers, drilled from within
the stope in the plane of the ore. The pattern, location of , ,
" convergencè points and measured closure are shown in Figure 4-16.
The holes were loaded with AN/FO and fired simultaneous1y. Three
rockbursts occurred 65 hours after the destress at the location
shown in Figure 4-16. The bursts occurred at a point where, . following the destress blast, little closure was measured. After
the rockburst greater closure wàs measured. Much damage resulted
from the bursts and to_date no further mining has been undertaken
in the area.
The next pillar destressed was the 1902 E crown pillar, located
in the 'A' zone on the 18 levél. As a result of an increase in
the microseismic activity and visual observation of the pillar it
was decided to destress' both the 1902 E crown pillar and 1802 E
sill pillar. The destress blast hole pattern is shown in
Figure 4-17. Holes were 45 mm diameter, drilled at 1.8 m
centers. The holes were loaded with AN/FO and traced with 't , 1 detonating cord, 1.5 m unstemmed collars were 1eft. Mi1lisecond
de1ays were used for initiation. Much micréseismic activity as
weIl as one rockburst occurred immediately after the destress
blast. Another burst followed a later production blast in the
pillar. No other incidents occurred as the pillar was
successfully mined out. During mining the pillar was observed to
be highly fractured, creating sorne drilling problems.
The most recent pillar destress was in 1604 E stope, located in
the 'A2' zone on the 15 level. The extraction ratio was much
l~ss than in the previous examples, as shown in Figure 4-18.
Based on previous experience and computer modeling it was felt
that the pillar would be burst prone once reduced to a 10 m
thickness. Spalling and 'fracturing were observed wlthin the
68
~
:!! ID C .. CD
.... 1 .... œ .... .... 0 1\)
m CJ (JJ -0
en '0 CD CD
~ II) ..
• 0 -.. CD ca .. (1) -3:: .. • CD ~ en c: en
". ()
"\ =r
• .. ~ ... ~
CD 01 ...
u
:
~ --S'ë) & ~~~BUètY~98Cl2 Tlmber:d
c..r~ ~'L-J' , 37 38 39 4 41 4 ~4~"-~~..J 1002 E DRI
1102 E STOPE
45mm Holes
~ , . . ~. pQW6ÉR'.:'IFi,CTO~ .;.p.e.6kQ/m •.•. '. .,." . ':.' .. '" (l.-(,,25 ·,6'lri3 . : .. " ~'. ': .. '. :. " : . -' . ': •. : .: 'Li' 0 'N"
.. .... :., .. '.:. :- :e{<PIO:Sf\f-e.s;;;-: Anfo ... ..... : .... ~ ... : :b~.IaY'.: :': < .... 25ms/hole
,. 12
10
• • ca C • • I! u •
mm 2
0
.. ;. ~"I .. t ).... ........ '( .. <' ......... "'''\ .... \'\ .... , .. .. : ......... -::-.... :.. '?'........ ::: .. .... ;~.~ .. ::f "$h .. :-:.<"::- :..... ...... ....:-
............. /$ .. :;;: ........ -:.-:- .. 1>-:......... .. .... .(':
, , .
Vay 3 J '1982
..\-: ..... ....... (>,\
" , -:. -:..... ... . ,
, .••.... .':OQ·l:Jb~e'Tru Llne
"
y _ EX1ensometer b Station
46
Closure From Extensometer Stations -
; .. " ::0 • ID •
k ~ ~ 10-1
.. 11 ,a 25
April
Il Il 23 30 • 13 20 27 M., June 1 9 82
'" 12
ta
& • c • • I! u •
mm 2
0
, \
-:! ::0 al
.. " Il 25 2
Apt Il
.. •
10-2
9 18 23 30 1 13 20 27
MI, June 198 2
1.
Il
l'
& 12
: 10
6 $ mm 2
o
~I . 1" J
.. ,t" ,. 25 2
~prlJ
10-3
-------• Il 23 30 6 '3 20 21
May June
".~ 1
..... CI
-~
ID --CD ... 1: ID 7t' c 0 :r
• -• , -ID C» -of
,1
::!! (Q c: ... CD .,.. 1 .... ...., ..
.... CO 0 fi)
m Cf) -0 '0 CD
'"tI
II) ... 0 CD (1) -... en CI)
en
~ ~~ .... '", .... ,t~ ....... : ~
tt'fACrOÂ' , '
~~'\,~>~~,,~,\, ,~~-- <,
~:: ... ~ç:::"~'" ~::- ...... tr ,~.. ... ':... .. ....
,~'.E~PJo,~f\tes, '.;. Antô 'l' "OëlOoa10{$ ,- 25
<~Oôtib là ,'t,unk f..' ~~ ,~~ .... ..;: .....
...... :: ... :.~ .. ~~' ..... .. .'" .. ~ '\ .... .~.. .. ..
( y ~"'", ~
.,/
: "= : .... " .. " . .,,' ........... • .. l " : ',',.' .. " ..
: ••• ~ ,," ~ .. " .. ,. • •• "" ". :. ' .. l " " " ..
, : ~ ....... :: ' , '! :... '. ': ' ; ',.' :,' . ': -: .-:: '.':: :. .: -:. ~ ::. '. : ' .. : ~ :: , : l .. ' : ,:,' ·1 8-0~: E: : Q .. ~:r~.E: ....... , .... : .. ~, . ..:".' t. .. " .......... 0" .' '..... ...0 ,",', ... "
'. ':. . '.' ..... ,:. 'I:t '1." .' •• , .... .' ••• '.: ... :. ::'J::'::" :.:.'; ::: :'; ': .: ,. .:: .:.: . ::, :.: . .' ;':: .:. ':,'. ~ ':: :.:'. ..,..... LF :, '" i __ .................... ,........... .. Jo "",,*,-,. ;:: A..-
" ... - - - .. ::. ".' 10 .... ....... .. ".. ...'. ...""
ni;&( 'l~1P':~E'tAEiNT" ~ ,"Ii,:::' ... ~ .. :'!':". ~ .. :': 'l'" ." .. .. • • • .." • .,., .. " .". ", , .. :" ~ ." .......... : " : .,.~. "
" 0' " '. "1 ... '.. :'0" .... ' ..... 0 •• : ~ • ". :" _::: •• ~ .:". ·E·· :C·· ·S'tOPE::.·· 's<';:"è: '7:. I:O':N : ........ : .... '._ .. .. .. . '.. ............. .. :........ .
, . • _. ~ "a •
0 ...... :.,
; ..
.. ,
., END
SECTION
.. ;: ... ~ ....... «1' .... ..-.. J'Y<;> ..... ; ... y ~ ,,<;- --- ------.,.,---- -------1802 E O,STOPE #-_ ... ~_ .. ---_ .........
.. ..,' ~1~..d +<"T· "~&l-4' ........ ".. ' .. d0 ~ ,':y 9. --~.'>~ " .... ".,",j., ~ ;;.-::.-:"" -_ >,,' ' ... ",~;.ti{lli. g",ç.J~.&il,- , "',, ". " " . '
" ~ ~ ~", ",.t;~,.,. ~ '" ' " -' ",' '45~ "DE .... ~ ~~"~'" ",,' '''',~~~. ~~..,..: ~..., , ~."" '.. -':::~-M.~",~Sn~.' ~.",. '!;"-<!J.~f,.,..J!lf(l" STRESS HaL ES
p N
y A
Extensometer StatIon
/'
c
•
,
(
c
· · . 1 \
, ' .' -. ' ..
, .' • .. . ,. "
Q) 1 •
-Q.;IO~ ....... \'
~ft.' • VI .'t4
.. ··~o-:" , , . .' , .. ..t" : ~.' -'-
'. , ., p ••
: . . ' '.
Figure 4-18: Long Sectlon A 1 Zone
after Makuch et al. 1987
•
stope but no unusual rnicroseismic,activity was noted. ln order to gain a better insight into ,the effects of ~estressing the stope -was heavilY:" instrumented in cooperation with the'-Canada Center for Mineral and Ènergy Technology' (CANMET) u~ing vibrating wire stress gauges, tape extensometer points, backfill pressure cell, blast vibration monitors, and a wave form analyzer'linked to the mine MP250 microseismic monitoring system, Figure 4-19. Disp1acernent discontinuity comPMter mode1ing suggested that rockbursting would occur in the ~entral portion of the pillar. The destress pattern used is shown in ,Figur~ 4-20. Holes were 44 mm diarneter, dri}led at 1.6 m centers from within the stope. The hales w~re angled into the footwall to avoid damage to the c,ernent
; ~ 1
~plug in the drift above. It was decided to destress no. 3 vein crown pillar, which was no longer econornic. This created a stress shadow for sub~~quent stoping of nos. 1 and 2 ,vein~, and avoided working through b1asted ground. It was also, necessary to blast a section o~ no. 1 vein to complete the destressing.
The destress blast was followed by only minor microseismic activity and no bursting. Measured ground vibration was within 20
percent of what was predicted. Backfill was not observed to carry any increased pressure. Few stress related problems were encountered as the no. "2 vein was rnined to the level suggesting "
that the blast was eff~ctive in Shiftcj~ng the stresses away.from the area: 'Sample closure and stress c nge data arising after the blast was compareato correspondi data predicted by a two
" dimensional, linear, finite element computer model, Table 4-2. The compari~on suggests that the blast resulted in a rnodulus
, \
reduction approaching 50 percent. One stress gaug~, located in the destress zone, recorded an increase in stièss suggesting tha~ the fracture plane did not extend between the destress ho les in
- ,
this area. Contrary to this, the nearest extensometer indicated average closure. The reliability of monitoring at this proximity to blasts is -questionable. The blasting may cause loca~
sloughing and incorrect closure m~~surem~nt~. Blastt~ration may cause 100sening of the wedge pinning the stress ~auges in place.
, 72
.~
C
~. ~
c
,e
""
'(
. -
, '
"
..J
\
\.
'.
--..,
o .
• Q. Q --0
~ W
~ fD ....
.. ... -.. Q
W
~, 1ft
"'"
"
....
~~------~---~~I •
Q
.. _L ____ !.
&----01 ':'"
........... ~. ,-,', 1 ", J" '0'.::: :.:. -':," .' ' . ," .. ...,.' .,'; ........ Q.
;/::::.::':.::::':-.i'.~ • : '~",t "I": ," •• : ,""".:.
...tJ'N', • 1
• , 1 \ , 1
c o :;:: Il U o ... " o ~ > e
1 1 1 1
1 . el 1 Z· , 1
~ 1 l' -' 1 L a. , ,
.. -• ~ ,. e o • c • ~ M W
, ,
-'
\ "
• .. l:I • • .. Go
. ~
Flqure 4-19: 1604 EW Stope',PlUar Instrumentation
/ a fter Scoble ~t al. 1987
73
_A'
'.
l,'
\
, '1-
Z : ' ..
w .. ,~ ..
w
•• CJ'
~-
c. ut C c(
o ..J
,ro .. . ,
'10. '
"
, ' . f
' .. • '. -1 ~ . . , .
.:.
, l';, 't.
.C;:) • . • Co.
'.
. r .'
OGlGl -cS C:O-J <z
VI W ...J o J:
...J
...J oC(
Figure 4-20; 1604 EW Stope Pillar .O~8tre88
. alter Scoble et al. 1987 (
..
w a.. o ten
w
M,
ê,.; CI': .x~ '11';; ~~ '0 o_
u a: o ~
u < u..
a: w Q ~ o a..
Table 4-2: Stress ehange ana Closure, 1~04E Pillar
StatioI) Measured Change 'Predicted Change 25% Modulus 50% Modulus Reduction Re~uction
Ex-2 -2.89 mm -0.7 mm -2.1 mm
Ex-4 -0.51 mm +0.1 mm +0.5 mm
E~-5 --1.41 mm -1.0 mm -1. 9 mm
Stress-3 -0.77 mPa -0.7 mPa .. -2.1 mPa
Stre"Ss-5 -5.50 mpa -1. 8 mPa -5.6' mPa \
Stress-6 +5.60 mPa +1. 8 mPa +7.6 mPa
after Scoble et al. 1987 .• -' ....
...
.. 1
c
. .. ..,.
) /-
M ,
C· " ~ --
fi'
"\ "- <" 0
t 75 ",
Il "
1
J
Core was recovered from the destressed pillars in holes drilled
between no's 3 and 4 and no's 15 and 16 destress holes. The exact
position of the diamond~rill holes between the destress holes i8 not known. Core analysi~ by the author revealed no difference ... in
the fracture density along the length of the core with the
exception of the collars. In order to asses microfracturing scanning electron (SEM) and opticai microscope studies were
conducted. Anastimosing microfractures were observed in calcite
crystals of one sample recovered from the destressed area, Figure
4-21' and 4-22. Quarhz crystals exhibited undulose extinction
while sorne mica crystals were highly deformed, both these
features are related to high stresses, but not shock stress. No microfractures were observed in samples from outside the
destressed zone. The samples outside the destress zone were' of a different mineralogical composition: less silicified and with no
large calcite crystals. It is thereforè no~ possible to confirm
whether the microfractures were due to the destressing or other o
stress related phenomena. No samples showed sflgns of fracture /
along grain boundaries which is a feature associated with blast
induced stresses. Descriptions of the thin section studies are
located in Appendix 1. "
Two Boxhole pillar destress operations ,have been carried out at
the mine. Boxhole piIIars are created between ore shoots in
stopes being mined by shrinkage methods. These pillars are found
to be the most burst prone structures at Campbell Red Lake Mine.
Destressing was conducted in piIIars showing signs of high stress
such as fracturing, spalling and squeeze on ore shoot structures.
On the 14 level in the 'F' zone the boxhole pillar.s were
located in a fault block area; it is believed that this resulted
in higher than usual stress concentrations. The pattern and
loading arrangement used is shown in Figure 4-23. There was no
bursting immediately following the destress blast, ,however, the ~-...",,,,,
followin week there was sorne bursting on the leveis above. Four
months l
area.
after
severe rockbursting throughout the stoping
observed to Qave increased significantly
Figure 4-23.
76
• ".
\'
'C
,
Figure 4-21: Thin Section Photograph of Anastimosing Microfractures in Calcite Crystals
\ --!igure 4-22: Scanning Electron Microscope Photograph of '-~-~~. Microfractures
~, 77 \
/ -
~ "TI co C .. tn
~ 1
ri,) ~
.... ~ ri,) .... 0 ~ ---Dl 0 .... )(
CO :x: 0 -CD
" -II) .., 0 (1) (1) -.., CD
DI (1) - (1) .. CD ... 3: III 1':' c: t)
:r CD
!!. ... 10 CD ...
o
~~ LJ9 1 ::le =.- , ~ 24
POWDER
.::,.,
"'~\:~\'{:\.'''''' ~ .. ,,\ .... ~~, ... ,," ..... ,\\",,~ ,~,~ .. -< " , ~ .. ,:: .. \"'~,\'.": ," \., ...... .... .. ,'.. ..
"
':::: ....... , .. ' .. ~'"\. ~ \: r '~)}jÇ:'---..-" ~\";"~\,~ _. ---- ~, \. .. .., .... \,'\" ,~'~ ... '~ ... \. ... ~' ....... ~~~\., ... , ... \ ..... , .. ' ~f"""" .. ~\" , .... 1 ..... ' .. \... -::,' + :::,~\ ... ~, ..... " .... '.. ...':: '\. .. "''':-''~~\~''\'''''' .. " ........
\' \ .... ~,~"'\, \. ","
t~
1 N A L
~ ~
<C
45mm Holes
:o"D 0 ~~~ 16 1,
ExplosIves - Anlo Delay -25ms/hole
SECTION Double Trunk Llne
"
..... .. .... ,,', '/
/'--- /14A: .17-,
.....t
Holes
Closure From Extensomete r Stat i,P n s • 5 DI C ..
• ti :&
mm 2
o
• 1
:. • li :&
mlll2
14 E
s •
14-2
• • < • ..
t~ p
• • ~
u
_--l--V /?
5 • 01 e • • 14 0 ti 3
mm 2
01
148·
01
• 5 DI .. ~ 41 14 C ..
• U l ~
.~ mm 21 è
0
5 • p 12 S • P 12
• 5
• ~ .. 14A ..
• •
l~ •
..c l • u !
'\ V 0
w
\)
c
(
«
<,
The second boxhole pillar blast was located on the 1500 level in
the 'G' zone. Single holes 3.6 m long, 44 mm diam~er were drilled into the centei of 18 pillars. ~The holes were loaded
, with Magnafrac 5000 cartridge explosives, a 0.6 m collar was
stemmed with çement grout cartridges. The holes were double
primed with Nonel caps and blasted simultaneously. A rockburst
occurr~d 2.5 hours after the blast in an inaccessible subdrift as shown in Figure 4-24. The burst a~peared to be located on an
andesite - altered rock contact. It is probable that the blast
was succe9sful ln destressing the pillars, however, stress
transfer caused a non-destressed pillar to burst. Vibrating wire Il
stre"ss gauges, located as shown on Figur~ 4-24, showed no changes
after the blast.
In addition to these relatively large destress blasts, the mine . ,
has also destressed_breas~ faces in narrow eut and fill stopes.
The technique has been to drill a single hole into the horizontal
face. The hole typically extends 1.2 to 1.5 m beyond the length
of the production round. The destr1ss hole is blasted along with
the round. The hypothesis is that of the original South African
work: an increased fracture zone would push high stresses further
ahead' of the face and provide cushioning for any burst that does
occur. The position of the destress hole was ~eft up to the
driller but is usually located in the upper center portion of the
face. Destressing has only been initiated when ~opes have shown
signs of high stress s'uch as bursting, increased microseismic
activity, spalling and stress cracks. It has further been
observed that the high stress concentrations are restricted to
specifie areas in the stope such as around geologic contacts, and /
near to raises, although reasons for local, stress perturbations
are not always obvi?us. Brief histories of two stopes in which
breast face destressing was conducted concludes this section.
The 1702 lE 'B' stope is located on the 17 level in the 'A' ore
zone. After one year with no mining activity stoping was re
initiated. In order to achieve high producti~n'the miner took . double lifts using vertical breasting. Very few rockbolts or ,
79
,
o·
CD r----.:-.-l. 10 10 -
altered rock
holes
o VWSt:4
hole.
\ ,
, (
Figure 4-24: Box Hole PlIIar Destres8 CRLM ( ~
80
1 .
, -
1
(
c
other support were installed. The stope became seismically
active abd high energy events were recorded by the rnicroseismic
monitoring system; eventually a burst occurred. In order to
bring the situation ufider control the stope was tight filled. \ ,
Horizontal breasting is now used with extensive rock support and
breast face destressing. The destress hales are drilled 4.8 m
while the round is 3 m, 32 mm hales are used in bath cases. Much A
seismic activity still occurs after blasting, however the energy
released is low magnitude and no more rockbursts have occurred.
This example has demonstrated effective stress control measures,
however, it is not possible to isolate the contribution of
destress~ng to the control measures.
The 1704 EW stope is located on the 17 level in the 'Al' ore
zone. The stope was experiencing high frequency and intensity
sei smic activity part{cularly after blasting. A number.of small
bursts occurred in the vicinity of a raise and at a dyke. The
problem was particularly bad where raise and dyke intersected. -
Destressing was initiated as a control measure. A 32 mm diameter
hole was drilled 1 to 1.8 m beyond the round depth and fired with
the round. Shorter rounds and heavy support were used in the . vicinity of the raise and dyke. A study conducted by the author,
Chapter 7, was unable to verify the effectiveness of destressing,
by this method, in this particular situation, however, as a safety
precaution the mine is continuing to destress this stope. (
81
/
o
5 PROBEX-l BOREHOLE DlLATOMETER
In order to complete this research project it was deemed necessary to be able to determine both in situ modulus of deformation and fracture density. After consideration of a
number of instruments, the Probex-l Borehole Dilatometer was
selected. This chapter discusses the operation, data reduction and problerns encountered with the instrument. Results and
Interpretation of data are presented in Chapter 8.
The limitations of laboratory derived Young's modulus for rock
mass evaluation are weIl recognized. Laboratory testing is
limited in representing the rock mass in question due to the size
of samples. Duplicating in situ environmental conditions is also
difficult in laboratory testing. Accurate measurement of modulus
of deformation is required for a complete understanding of r0ck
mass behavior. Coupled with the need for accurate mOduli values
is the need to detect and orientate fractures at remote
locations.
Several instruments have recently been developed for the in situ
measurement of rock mass modulus in diamond drill holes. The
instruments measure the response of the rock to an applied load.
If the load is applied in a uniform, radial, manner then the
instrument is termed a dilatometer or pressuremeter. It is
generally accepted that a dilatometer is for measurement in rock
'while a pressuremeter is for measurement in soils or soft rock.
An advantage of borehole dilatometers is their ability to obtain
measurements in locations accessible only by drill hole.
Continuous in situ measurements along the length of a borehole
allow rapid testing through a large volume of rock. The instrument is particularly valuable where core recovery i8 poer,
provided that the boreh~le walls are not too irregular.
Dilatometer rneasurements are preferable to laboratory tests as
the volume of rock tested is greater and dilatometer tests can be
conducted in moderately fractu,red and jointed rocks.
82
c.
c
oilatometers can determine variation in deformation modulus as a
reBult of changes in lithology, 'joint frequency and areas of 1
weakness such as fractured and weatherea zones. Dilatometers have been used ta outlined zones of excavation damage, Koopmans
nnd Hughes (1985, 1984), measure the deformability of dam site
foundations, Rocha et al. (1970), measure short term creep in t
potash, Ladanyi and Gill (1983), and evaluate stability in .
underground hard rock mines Scoble et al. (1988).
5.1 DESCRIPTION-OF INSTRUMENT
The Roctést Probex-1 is a flexible dilatometer with a
cylindrical, radially expandable adiprene membrane. A schematic
diagram of the instrument is shawn in Figure 5-1. Pressurilàtion
is achieved by injecting fluid from the water cylinder into the
membrane cavity. The downstream piston in the water cylinder iB
actuate? by the hydraulic pump situated outside the borehole.
The piston is instrumente~with a linear variable displacment transducer (LVDT) to measure injected volume; this arrangement
eliminates errors caused by the parasitic expansion of the tubing and pumping system. The LVDT digital read-out unit is located on
surfàce. An initdal calibration determines the digital units per
injected volu~. Pressure is measured from the dial gauge
situated on the hydraulic pump. This measurement method requires . correction factors for the inert~a of the expanding m~mbrane and
dilation / compression (stiffness) uf the testing system. The
membrane inertia proves to be negligible when testing in ha rd
rocks. Other dilatometer systems incorporate LVDTs within the
expanding membrane to measure volume expansion directly. This
arrangement has the advantage of not requiring a correction Jor
system stiffness, and of being able to detect anisotropy_within the rock rnass.
~he Probex-l is 3 m long and weighs 40 kg. Operation of the
instrument requires two people." Both drill casing and wire 1ine " ,
rnethods have been us@d fo~.insertion of the instrument into the
83
o
Inflation Une
Deflation Line _--f--t-t---t Pressure Gauge (1 nflation Circuit)
Pump
LVDT ~~.uL1
READOUT UNIT
1 Dilatable Membrane
Il
Pressure Gauge (Deflation Circuit)
MEASURING MODULE
Upstream Piston (Oil)
DUAL-ACTION HYDRAULIC MODULE
Dual Piston Rod
i!~h- Downstream Piston (Water)
a: llJ
h ~_ Cylinder ~ ~
EXPANOABLE DILATOMETER PROBE
Steel Core , (Mounting for DiJatometer Probe)
Saturation Plug __ -"("
Figure 5-1: .Schematlc Diagram Of Probex-1
.p 1
1
after ROCTEST (1986)
84
c
c 1 1
hole. The instrument attaches to !BX' size'drill casing. An
attaehmént for wire line operation was constructed at McGill University, Figure-S-2. Figure 5-3 shows the instrument at an
underground test site. ~
"
5'.2 DEFORMA'l'ION MODULUS DE'l'ERMINATION
, The standard analysis of dilatometer data pssum~s the rock mass
to be a linear elastic, homogeneous and isotropie continuum.' The
effeets of anisotropy on modulus of deformation'can, however, be
substantial. Amadei (1985) has published a theoretical solution . "-for determination of deformation modulus in anisotropie rock.
The effect of in situ stresses on dilatometer results is also of
&oncern, Ladanyi (1975) diseusses this problem. The ISRM (1986)
suggest a method of analysis for regular jointed rock masses and
also recommend adjusting fo~ scale. In most situations the,
standard analysis is still used as in situ stresses and
discontinuity properties are very eomplex or unknown. rhe cv
limitations of the testing method and method analysis must be
recognizedi results must not be confused with other modulus
values. In this text, results are referred to as dilatometer moduli of deformation.
If the loaded length of borehole is large eompared to the
diameter, the·stress regime is considered to be plane strain.
Assuming an elastic medium the stress at any point is given by:
~
where
6rr = Pi R2 /r2 for compression
for tension e = -Pi R2/r2 "
ôrr= radial stress
r = radius of borehole d' 'b ('d Pi = pressure, uniformly l.strl. ute
R = distance from center of borehole
85
-,
)
. ,
National Library of Canada,
canadian Theses Service
NOTICE
o
THE cjU~LITY OF THIS MICROFICHE 15 HEAVILY DEPENDEN" UPON THE OUAL~TY OF THE THESIS SUBMITTeo' FOR MICROYILMING.
UNFORTUNATELY THE COLOUREO ILLUSTRATIONS OF THIS THESIS CAN ONLY YIELD DIFFERENT TONES OF GREY.
'\
l' 1 :
( .
o 1
"
1 :
Bibliothlque nationale 4u Ca~ada~
Service des th'ses canadienne. "'"
--~ AVIS
"
o
LA QUALITE DE' CE"TTE MICROFICHE bEPEND, GRANDEMENT DE LA OUALIT.E DE LA THESE SOUMISE AU MICROFILMAGE.
"
• MALHEUREUSEMENT, LES DIFFERENTES
, ,
ILLUSTRATIONS EN COULEURS DE CETTE 0
TBESE NE PEUVENT DONNER OUE- DES TEINTES DE GRIS.
• <>
"
u •
. -.
•
..
a-
~\ \
• ,
1
Figure 5-2: Wireline Attachment for Probex-l Dilatometer
Figure 5-3: probex-l At Underground Test Site
86
c
,
c
(
·. --ri~·- --------r------------__ -r '
p
The volume involved in a single test'~pènds ~p.n. the pressure '" used. Using the above equations it is possible to estimate this ,
volume. Theoretically the dilatometer induces.stresse~ in the rock out to an infini te distance (r). The minimum stress, beyond which the rock is no longer influencing test results, is not,as yet defined. Exam~nation of"a pressure volume curve indicates
that the rock behaves in a linear fashion after being str€ssedbeyond 10 MPa. If 10 MPa is considered to be the stress below .
t which the rbck ~s no longer influential of test results, then the volume of rock 'affecting the test is calculated to be 0.005 m3 .
The dilatometer measures the diametral deformation of a borehole subjected to a uniformly distributéd radial pressure. Using ..,
- thick wall cylinder theory and assuming an infin~te elastic medium, the rock modulus can be deteFmined.
The deformation of the diameter of the borehole is given by:
or dr/r = (l+v) dP/E
E = (l+v) r dP/dr
where v = Poisson's ratio
E = modulus of elasticity P = pressure in probe
\
dr = change in radius of borehole
, <
dP = change in probe pressure between
= Pb - Pa
Pa = pressure at first measurement
Pb = pressure at second measurement
readings
When expressed in terms of volume instead of radius, the equàtion becomè-s:
E • 2(1+v) V dP/dV
87
J
.. . in volume between readings
, "",
• 'h.
' .
V = volume of cavity when dP and dV are 'mèasured '- = Vo + Vm
Vo = initial volumë of deflated probe Vm = mean' additional volume injected
= (Va + Vb) / 2 Va = volumè at first m~asurement
-
Vb = volume at second measurément
" The pressure and volume change must be corrected for membrane inertia and syst'Em.-dilation / compression. The equation becomes:
tE = 2, (1 + v) V
dV - c
(dP-dP i)
.. where dP· = change in pre$sure due to inertia &f membrane
1 ~
cOFresponding to increment dP c = volume correction factor due to system __ dilation /
compression.
\.
5.2.1 DlLATOMETER CALIBRATION FACTORS
• The Probex-~must be corrected for membrand inertia and . . dilation. Membrane inertia is established by inflating
system the ~
unconfined probe and plotting a pressure versus volume curve from . , which the correction factor is 09tained, Roctest (1986). During , tests it was noted that sorne of the membranes did not inflate symmetrically due·to variable ,embrane rubber thicknes~ (12-16 mm).~ Modulus deterrnination is \ot affected by this. Best fit curve for the membrane inertia data was of the form:
y =, m Xn
88
Je
(
f,
- where y = pressure X = volume m and n are dependent upon -the membrane used.
Regr~ssion coefficients (r2) of 0.99 were obtained. The maximum pressure required to overcome membrane inertia was 1~2 MPa, this
I(Y 1" __
'is insignificant when testing in ha rd rook. Repetition of the inertia correction curve was good, however, tpe maximum pressure
~; , 1
required to overcome me~rane ,inertia decr~ases slowly with us,e.
The volume correction is determined by inflating the probe in a calibration cylinder of known stiffness and plotting a pressure versus volume curve. The slope of the lipear portion of the curve is the stiffness of the combined 'testing system plus calibration cylinder. Compression occurs in the inflating water and rubper membrane while dilation occurs in the steel core, reaervoir casing and piston. The volume correction calibration .. factor 'c' is determined by the following equation:
\ c = a - b \
\ where a = inverse of testing system and cylinder stiffne~s, ,
b = inverse of calibration cylinder stiffness
b = 2Vc (r+t (l+vct)
. '
--.. ftJ
where Vc = inside volume.qf cylirder'affeeted by test .:>
r = inside radius of test cylinder •
t = cylinder thickness~ Vc = Poisson's ratio for cylinder material
.1 Ec = Young' s modulus for cylinder material
The factor 'e' is the inverse of the testing system stiffness. Stiffness, in this text, is defined as the stress deformation ratio (kPa/cc)~ The accuracy of the factor c is crucial to the
89
,\ ,
1
..
o
, modulus determination. A difference of- .00001 cc/kPa can significantly alter results.
)
A typical calibration curve is shown in Figure 5-4. Data
analysis reveals the slope to be hi9?ly variable, depending on the portion of the curve over which the slope is taken. At high pressures, i.e. over' 25 MPa, there is an increase in slope and. non linearity of the curve. This is due to the non linear , deformation of the instrument. Results are-al~~ dependent on the
/
method and accuracy of the instr~~ent readings. Table 5-1
presents a listing of calibration results obtained for three
membranes. Even when tests were performed under identical conditions, the reproducibility was not good. All data was . ( ,
analyzed over a 13 to 26 MPa pressure interval by'linear " \
regression. During calibration tests the system was slow to
stabilize, at high pressures stability could take ~s long as 10
minutes. This is due to time depen~ent system dilation and heat loss of the compressed fluid (this component is small). The
slope of the curve decreases with time which resflts in a higher_ calibration factor. Our procedure h~ been to ufe an average of
ma~y results ~btained after readings stabilize. ~
Other authors have reported leakage problems with their
respective ditatometer systems; this would create reading
instability. Leakage has occurred in the McGill instrument,
although, it has not caused any significant problem. Ruman error
in reading the pressure gauge i8 believed to be the great~st source of error. A pressure transducer and digital read-out are
recommended to overcome this Pfoblem (now availablé through RocTest).' The calibration curve waa observed to ahift to the
',---left when testèd underground.' Thi~ did not appear to affect the ,
calibration factor. ft
The calibration factor 'a' ia dependent upon the diameter of the
testing cylinder. Membrane end effects and thinning are
influenced by the hole diaméter. Calibration factors should be determined over a range of different sized cylindersl the value
\ \ \
\ \
90
c
. o
c
~f('f'.
28
21
2.
22
20 ..... ~ ,- t8 Go :1 ..... ta • ~
t. :t
• • • t2 .. Q.
- 10
1
a
• 2
0
y)
a Clonflned calibration
• 'In .It u te .'t"
~ -. J 1
!-~
Q InJected volume (cc)
..
\ \
200
figure 5-4:Typlcal Pressure Versus Volume Test Data
• ~ :1
• • • .. . Q.
" • --ca. ca. •
\ ,
'C curv"e curv. urve 2 curv. 4 1 3
Figure 5-5: Idealized
Pressure-Volume
Deformation Curve
/ InJ.ct.d volu ...
a fter Scoble et al. (1988)
91
..
.. . TABLE 5-1: CONFINED CALIBRA~ION..RESULTS
Q PROBE 1 LOCATION DATE CALIBRATION COMMENTS
FACTOR 'a' (cc/kPa)
114 McGill 8/12/86 .00114 first calibration 114 mine 8/12/86 .00118 114 mine 8/13/86 .00117 under ground 114 mine 8/18/86 .00119 under ground 114 mine 8/18/86 .00117 under ground 114 mine 9/14/86 .00119 under ground 114 mine 9/14/86 .00121 under ground 114 mine 9/14/~,6, . ~0119
125 RocTest 5/2:t/8~6 .0012ii , first calibration
125 Roc Test 5/21/86 .00118 125 RocTest 5/21/86 .00123 125 Roc Test 5/21/86 .00122 125 RocTest 5/21/86 .00125 125 RocTest 5/21/86 .00115 125 RocTest 6/2/86 .00119
3 125 RocTést 6/2/86 .00121 125 RocTest 6/3/86 .00120 c~~~
125 RocTest 6/3/86 .00120 ~
125 RocTest 6/3/86 . 00120 .. ">
125 RocTest 6/3/86 .00122 ~ ... 1.
125 RopTest 6/3/86 .00118 125 McGill Il}4/86 .001228 125 McGill 11/4/86 .001237 t • .. , 125 McGill 1/30/87 .001377 readings after 5 mfn>': .~
125 McGill 1/30/87 .001360 rea(iings after 5 min. ": 125 McGill 1/30/87 .001273 readings after 10 min.':' 125 McGill 2/13/87 .001328 readings after 6 min 125 McGill 2/13/87 .001284 readings after 1 min. 125 McGill 2/13/87 .001281 readings after 1 min. 125 McGill 2/13/87 .001280 readings after 1 min.
(j~i 125 McGill 2/13/87 .001273 readings after 3 min.
-~ 125 McGill 2/13/87 .001296 readings after 3 min. ";o!>l .. ? 125 McGill .f 2/13/87 .. ,0,01310 readings after 6 min.
125 mine 8/1/87 .001197 under ground l.('~. 125 mine ' 8/1/87 .001273 under ground
1. 125 mine 8/1/87 .00l-'332 under ground ~<.".~-'"., t
"
-107 McGill 2/15/87 .001192 first calibration 107 McGill 2/15/87 • 001133 readings after 1 min • 107 McGill 2/15/87 .001137 readings after 1 min.
0 .107 McGill 2/15/87 .001146 readings after 1 min.
107 McGill 2/15/87 .001153 readings after 5 min. 107 McGill 2/15/87 .001162 readings after 5 min. 107 McGill 2/15/87 .001161 readings after 5 min.
'\. 92
'1\
c
r
c
,
~
-, ~of 'a' u~ed in determination of modulus should correspond to ~he # 'i f... .1
diameter at the test location. --(
~"
A correction, to account for pressure head,' is required when 1 • 'II
testing in vertical holes. A pressure equivalent to- the static head ~f the hydraulic fluid lipe ~s added to the gauge ~
pressure. The value is insignificant from 0 to 8 m below the elevation of the/gauge.
~.2.2 DBTIRNŒNATION or BOLI DIAMBTBR .;;'
It is possible to determine the approximate size of the borehole, at the pofht of testing, by observing the volume at which contact
... '. ia made with the borehole wall. Wall contact i5 indicated by'a pressure increase. Some initial pressure increase is expected due to membrane inertia." Diameter i5 determined by the following formula.
;., -. where
o = 2 V (V 0 + Ya)
11" l
o = diameter of hole -:, ...
.Vo = initial volume of probe Va = volume added to contact borehole wall
l = probe length
The equation àssumes equal inflation along the length of the ~ membrane. It ignores end tapering and thi~ning of membrane material. A second method to determine volume is to mea~ure the diameter and volume of the unconfined membrane as it i5 inflated. The borehole diameter is determined by mat ching the volume at which wall contact is made with the unconfined volume and diameter. Detecting the exact volume at which the probe seats itself is difficult. With both t~chniqu~s, the calculated diameter w~s found tq exceed the actual diameter.
J
o
/
5.2.3 PROBBX-l TBSTING CONSIDERATIONS
The membrane is r~ted to a maximum ing pressure of 20.7 MPa in a 82.5 mm'borehole and 30.0 MPa n a 76.2 mm borehole. These a're values at' which membrane rupture as occurred, intermediate
"values are not known. Maximum pressu es well below these values were used to avoid rupture. Hole diameter was determined using the techniques discussed in the previous section. Rupture has resulted from material flaws not linked to excessive pressure.
Readings taken during testing, as with calibration, are slow to
stabilize. This may be due to leakage, time dependent 'system dilation or time depen'dé,nt rock dilation (creep). The technique
used during testing was to allow a con~tant stabilizati~ime for each measurement; ~aiting for complete.stability proved too
• time consuming. The method suggested by the manufacturer is to
take readings at one minute intervals until the volume change is
within 5 percent of the initial reading. This method,Js only suitable if the percent age difference is always the iame. The
following two cases would yield different results. Case 1:
volume reading at pressure 1 is taken after a 5 percent
difference has been reached, reading at pre~sure 2 is taken after
a 2 percent difference has been reached. Case 2: votume reading
at pressure one and 2 are taken after a 2 perèent difference has been reached. Koopman~ and Hughes (1984) have improved the data
collection on\.their dilatometer by using an electric pump and l"
automated datq acquisition system.
The testing proce8ur~ used was to inflate the membran~the desired position to seat the probe. The pressure was then
, released and allowed to drpp to just above the seating pressQre.
A m~nimum of three loading cycles were then rune Unloading .. cycles ~ere attempted, however, they proved to be extremely
\ difficult with the pump arrangement used.
Data was'collected along the entire length of all ,bore_holes.
D:e to system r~liabil~d experimental error, dilatomete~
94
c
( • 0
J • , .
\.
c
deforPlation modulus value should not be a'ssiqned to q.iscrete test location. Values should b~'averaged and assigned to àn area of the rock mass.
Fig~re 5-4 shows a typfcal hard rock load versus deformation curve. This curve includes both thé rock and testing system stiffness components. As with the calibration curve" thé rock plus testing system slope is not linear. Stiffness depends on
" the segment of the curve over which the slope is analyzed. The slope increases at higher pressure values.
When testing in high modulus rock a number of,additional problems arise due to the small increase in rock volume as compared to the change in testing instrument volume. The sensitivity of results to experimental error is greatly increased. In sorne cases the stiffness of the rock plus testing system exceeded the stiffness of the testing system alone ie. rock stiffness exceeded testing system stiffness. In this case meanirigless negative values for deformation modulus resulted •
Figure 5-5 schemati~ally illustrates the problem of testing in high modulus rock. Where rock plus system stiffness, curve l, exceeds system stiffness alone, curve 2,'negative modulus values result~ Anomalously high modulus values occur when the rock Q
plus system stiffness apprOach that of the system alone, curve 3. Curve 4 represents acceptable results. If the rock plus system stiffness falls between curve"l and 3 erroneous results will oeeur. Analysis of the data indicates that if the rock plus testing system stiffness approaches within 4.5 percent of the testing system stiffness then erroneous results will occur. This is due to the sensitivity of testing at high pre~sure and low deformation. When testing in high mo~ulus rock a large portion of the volume change is due to system dilation / compression, and only a srnall portion to rock dilation. As the percent age of dilation attributed to the testi~ system increases, the ~uracy
.~
of·rneasurement deereases.
95
o \
1
o
l 2
\
Where results arg negative, or erroneously high, reasonable • •
values can be obtained by analyzing the slope of êhe rock plus <
system curve over a lower pressure interval. This results in a . ~ -
lowey stiffness value, h~wevër, the volume of rock con~idered in the kest is reduced. Analysis over a 13 to 20 MPa pressute interval was the standard practice used during this research.
r.- •• •
A second approach t6 allevtati~g the problems of testing in high modulus rock is. to increase the testing systemstiffness. This ,
can be done by replacing sorne e~the water ~ith a more stiff material such as steel balls. At 30 MPa water is compressed approximately 1 percent, the total water capacity of the Probex-l is 1 liter. Reducing the total water volume by r~placing it with , steel balls will significantly increase the system stiffness.
~
Reducing the wat~r volume, however,. would reduce the' capability te test in oversized bo~eholes.
The validity of dilatometer testing in high modulus rocks ~ppears questionable. Kaneshiro et al. (1987) suggest an upper limit of 41 GPa with the Probex-l. The author, however, has meaaured in situ dilatometer deformation moduli closely comparable to laboratory values of Young' s modulus in rock, up to, .80 GPa. .1
.. 5.2.4 SBNSITIVXTY AND RBP&ATABILI~Y OF
MODULUS or DErORMATIQN RESULTS ..
" The ability of instrument~ion to yield repea~ble and accurate results is of utmost impo tanè~. In order to asses the Probex-1 performance a number of st di~ were conducted. Repeatability
./
tests were Cfnducted in the calibration cylinder as weIl as in ha rd rock boreholes. A study to evaluate instrument gauge
,/
sensitivity was also undertake~ In situ results were compared to labo rat ory derived modulus values. /\
. . Tests conducted in the calibration cylinder should reflect the repeatability of the Probex-l. The cylinder ia of known ~terial
~
.96
J'
- .--
c -/
c
. . ,
___ 1
properties and is not affected by repeated testing. As is , discuBse~ in the section on calibration, results of repeated , tes~s under standard laboratory conditions, have been variabl~.
~ '" " Repeated tests within a rock borehole wére found to be variable. In borehdles with few fractu~es, subsequent test generally yield higher modulus vâlues. Increases are due to closiqg of pores and cracks within the test area. In highly fractured boreholes, subsequent tests generally yield reduced modulus values. -The reduction is assu~edJto result from hole degr~dation c~used by
,c the first measurement. Reductions from 0 to 20 GPa occurred • . Increases have been from 0 to beyond the point of measur,ement, ie. \he ~tiffness of the rock plus testing system curve exceeded that of the testing system alone.
A sensitivity analysis was conducted on the accuracy of instrument readings. Specifie results ar~ dependent upon the data used in the analysis, however, trends are similar: ,
<l..
altering the recorded volume inj~ted, at a single reading by as little as 0.05 ec significantly alters results.
, altering the recorded pressure, at a single reading, by as little as 50 kpa significantly alters results. altering _aIl recorded volume readings or all recorded pressure readings by 3% significantly alters results. sensitivity to reading accuracy is g~eater at higher .. ..... pressures.
al~ering aIl recorded pressures or volumes by an equal qu~ntity does not affect the results.
A comparison of dilatometer derived modulus and laboratory , derived modulus frdm uniaxial loading was under,aken. It should
'\-be noteq that the dilatometer y1elds modulus r~sult$ for the plane normal to the bore hole axis, while core samples measure modulus parallel to the borehole axis. Core samples are from competent sections of rock while dilatometer tests are conducted over fractured ground. As discussed earlier, modulus values are
97
,(j' •
..;.
o
dependent on the orien~ation of the test with respect to an~sotropy. Th~ ,results, Table 5-2, show htgh ,variability. oilatometer deformation values both lower and higher than laboratory derived'deformation modulus~~eurred. In the case of high modulus rock~, the dilatometer gave the lower value, where as with low ~odulus rocks the di1atometer returns a higher value.
Hustrulid and Schrauf (1979) report a 35 to 40 percent reduction from laboratory values using a Colorado School Of Mines dilatometer. Kaneshiro et al. (1987) report Probex-l results as being only 40 percent of laboratory values. Correction factors have been utilized by sorne researchers, while others have consi-_ dered dilatometer d~rived modulus a lower bound value for the rock mass.
5 .3 IMPRESSION PACKING
Impression packing is a means ,whereby an impres~ion is made of th(! surface of a borehole wall. The impressi~n records' fractures and other indentation or protruding features in ,the borehole. The tec~nique with the Probex-l is to wrap the membrane in a material that will record the borehole features after it has inflated at the desired location. A number of different materials have been tested to deter~ine the most suit able for impression recording.
;
-.a been
The impression technique has fts·limitations although at\present not weIL defined. Simple experiments indicated that a c~ck that could be detected by a fingernail (approximately 0.1 mm) would , -create an-impression. Hairline fractures not detectable by
fingernail would not create an impression. Fracture orientation,
packing pressure and the effects of in situ stresses ,have yet to - -- J</
be fully explored. Reductions in modulus have been~detected
where no impressions wereQrecorded. It is assumed that fractures
too small to be resolved by the impression technique yet significant enough to alter modulus existed.
98
c
c
c
. -'
Table 5-2: Probex-l versus Laboratory Deformati~n Modulus
Modulus (GPa) Sample Probex-l laboratory
1 69 83.0
2 50 81. 3
3 18 7.4
4 22 21.1
5 33 20.7
6 36 17 .5
7 29 22.0
8 19 12.5
9 47 26.7
Koopmans and Hughes (1985) have reported similar findings. A
~ 1-imitation may weIl be the result of the material used to record ~ \
the impression. Vinyl cOding tape or electrical tape proved to be the most suitab1e. Red was found to be the best color for working in an underground environment. Other impression materials tried made poor impressions, were too thick, or were
~
not durable enough.
The procedure to mak~ an impression is as follows:
tif; wrap the membrane _~ith tape, overlap to ensure complete coverage. insert the probe ta the desired position within the borehole
and inflate. \
mark a reference 1ine on the drill casing string to identify top of hole. - remove probe from boreho1e.
o
trace impressionsonto plastic film, marking the reference line.
·99
(
--
o
. ,
Figure 5-6: Tracing Fracture Impressions From Dilatometer ......
,t
• J
•
100
c
• •
•
c
y \
( F!gure 5-7: Fracture Impression Maps
b
101
o
o
N ,0
N
h
102
Flgu r. 5-8: Incllned Ol.'è ontlnulty
Plane eutt'n g Drill. Hoi.
,Figure 5,-9: Impression Map From
Drill Hole
c
c
,1 !
c
..
Figure 5-6 shows the impressions being traced from the probe onto . ~lastic film. The resulting fracture maps are illustrated in Figure 5-7. A reference line is only possible when using drill casing insertion. It is still possible to orient the core, withçut reference line, if an impression Qt a known feature is
~~" -
recorded.
If the drill cote is of sufficient integrity it is .possible to physically align the drill core in the hole direction with co~rect rotatiQn. Discontinuity directions can be measured
1 .
directly from the core with suitable instruments, ie. compass-inclinometer.
It is also possible to obtain discontinuity orientations from borehole impressions. Figure 5-8 is an inclined discontinuity plane passing through a section of drill hole. Figure 5-9 shows the impression map created from this borehole. The reference line (Lr) represents the apex of the borehole. The line of maximum dip .(Ld) is also shown. In the case of a horizontal or vertical hole, determination of discontinuity orientation is found by the following equations:
dlp = arctan h/D d
dip direction = e + direction of Lr
where h ='distance between sinusoidal peaks D = borehole diameter e = angular distance between Lr and Ld
_ ,- Ii
.. If the drill hole is inclined the,above method yields the
, apparent dip. Sterographic projections are the most convenient way of correction. Figure 5-10 is an equal angle net with an example solution. The borehole (and reference line) trend 315 degrees and dip 30 degrees from horizontal. From the 76 mm diameter hole impressio~, the distance between sinusoidal peaks, h, is measured to be 64 mm resulting in a 40 degree
103
•
..
,
o
\
" , ,6'0
" l , , , , ~
Nd
FI~re 6-10: E,qual Angle Stereograph Wlth Example Solution
r r 104
~" \
/ - ---- ... --..'
apparent dip. The angular distance between Lr and Ld (e) is 60
degrees.
-_\~ The-borehole (B.H.) is brought to the vertical by a 60 dégree ~ /
c
/'
(
-' auxiliary rotation along the east-west diameter to B.H. ' • On the stereo net perimeter 60 degrees (9) is counted of from the 315
( degree mark. The apparent dip is counted-off along the east-west diameter to establish point D'. The normal to D' is found, ND'. The pole ND is determined by recovering the original 60 degree auxiliary rotation along a smaii circie. From this the
..
, . .::. -discontinuity is determined to he: dip 49 degrees, dip direction :qJ
86 degrees.
An area of concern was the potentiai damage to the impression
recording material dur~' , probe insertion and retraction. A tape wrapped membrane was ru p and down a horizontal three meter
~ long borehole three times. Very few marks were observed. AlI
the marks produced were either parallel to the probe axis or along the tape edges. Most marks occurred along the botto~ of the membrane where it made contact with the borehole. Marks created by probe movement are straight edged while those due to ' fractures are irregular. Movement marks and those caused py inflation over'rock debris are indentations while fracture
, l
impressions are extrusions. Impressions due to rock cuttings may
~bscure the fr~cture pattern, it 15 therefore "recommended that -
holes be flushed weJl prior to testing. Non fracture markings ~
are readily identitiable and ornitted f;om analysis.
1
A number of impression packing repetit~ons were conducted. Figure 5-7a and 5-7b are su ch repetitiorts. The pa~king pressures
\ were 20 MPa for Figure 5-7a and 26 MPa f6r Figure 5-7b, the
, higher pressure used yields a more detailed impression. Reproducablity is considered good.
\ J
'" 105
.. "
,
o "
o
)
5.4 SOMMARX AND CONCLUSIONS ON TBB PROB&X-l DlLATONITSR
The Probex-l borehole dilatometer is capable of yielding a relative value of rock mass deformability. Magnitudes and locations of deformation changes are detectable by the instrument. Care and attention on the part of the user ls ~~ired for effective use. The instrument is wel~ suited for in situ investigations of blast related fracturation and modulu8
, ' reductions. particular noteworthy points are as follows:
> the accuracy and repeatability limitations of the instrument indicate that modulus values assigned to the rock mass should be based on the average of as man y tests as possible. This
-also applies to the derivation of the probe ~alibration factors. to reduce instrument reacling errors a pressure transducer with digital read-out is recommended. modulus values should be calculated using as high a pressure as possible to incorporate the greatest volume of rock in the
test. instrument readings should be taken after stabilit~~as
-J been reached. Where this is not practical due to time constraints, then readings should be taken after a fixed time
, interval. , "
the Probex-l can be used effectively as an impression packer, this is useful to aid in evaluating results in fractured ~
and anisotropie ground. - -a more complete study is required to assess the applicability
of dilatometers for testing in high modulus rock.
106
c
c
c
6 MICROSJlISMIC MONITORING
Microseismic monitoring has become an accepted tool for ground control purposes. Structural instability can be detected and located, as can areas of high stress. Discontinuities can be located and delineated. The method is based on the detection and analysis of mi~roseismic energy waves that originated in, and travel through the rock mass.
The origin of naturally occurring microseismic energy is unknown . . Hardy (1987), attributes it to the sudden release of strain energy which generates an elastic stress w~ve. These r~actions are due to the ground trying to achieve stability under changing stress conditions. Possible causes are divided,into micro, macro and mega levels of emissions. At the micro level it is related to dislocations. At the macro level micro sei smic activity is related to twining, grain boundary movement and the propagation of microfractures through and between mineraI grains. At the
mega level it is r~ated to fracturing and failure of large volumes of rock or m6tion between two structural units. McGarr (1984) reports that there are no systematic physical differences , /"
between the microseismic waves generated by rockbursts and those generated by small earthquakes. Microseismic energy can also be imparted into the rock by a physical means such as blasting or sudden impact.
This chapter discusses the microseismic wave form and the equipment used at Campbell Red Lake Mine to monitor microseismic emissi~ns. The application of this equipment to evaluate ' destress blasts is also discussed.
6. 1 '1'SB NICROSIISNIC 'tIAVB l'ORM
Microseismic wave motion is recorded as either a displacement variation, a velocity variation or acceleration variation with respect to time or distance. The most common monitoring systems
\
107
"
~
o
L
record either velocity or acceleration with respect to time.
An idealized microseismic event, recorded as a displacement time • history is shown in Figure 6-1. The wave can ~e characterized by a number of values including amplitude, rise time, duration,
~article velocity and acceleration. The threshold limit is the amplitude above which the wave must pass in order to be considered for further processing. It is an arbitrarily selected value. The threshold is usually set at 2 to 3 times the background or cultural noise'level. D~sp~acement, velocityand
acceleration are aIl numerically interrelated as shown bY,the
" following expressions.
dU dU
u = Ü = dt dt
u.= displacement
u = velocity
Ü = acceleration
.. ,
As with blast vibration monitoring, the wave form can be considered as a sinusoidal approximation. The advantages and limitations are presented by Dowding (1985). The simplification
allows the following for;!lulae to be develope'd.
il = 2 uf
ü = 2 uf
f = freqJ,lency
Microseismic events generate both shear (5) and compression (P)
waves, known as body waves, which travel through the rock masse Raleigh and Love waves, known as surface wa~es, are created when
body waves reach a surface. The P wave has the greatest
velocity, while most of the enérgy iB associated with the S wave
or Rayleigh wave if at a surfate.
108
("
c 0 .. CIl t ... ~ '0 .. C
• > •
\
.L ... . . . ... .... . . ... ... ~ ...... . ... .. ~~-+------+--~-- -....
c • • ! e • .. - a. .. 0 • • • • • • " - - ~ ... .. .. .II: a. • • e . a. •
'0 C ::. 0 .. Q • .II: • U CIl 0 s:. c
~ --4
~ ) .II: • • • '" " Q. :2
" = • 0 Q. 0 • .. oC • .II:
e • s:. • • \ • ... • oC
\ Q. .. ,
Figure 8-1: Mlcroaelamlc Event
~. modlfl.d aU.r Blake 1882
C
•
)- '
o
It is possible to represent a microseismic signal in the time d~ain, as in Figure 6-1, or in the frequency ajmain. Conversion to a frequency domain is achieved by numerical methods such as Fourier transforms or pseudovelocity response spectra, Dowding (1985) .
- .. Frequency spectra of seismic events at their source are unknown. Laboratory studies suggest they are fairly broad band, from 300 to 300,000 Hz, Hardy (1982). Small events, such as the formation of microcracks, generate high frequency waves, larg events produce lower frequencies. Lower frequencies contain a greater #
portion of the seismic energy.
Small, high frequency events have1pulse durations of . milliseconds, larger events may last for more than ten seconds. The duration is a function of distance. Frequency separation and wave type separation occur with distance. Amplitude ls a function of size of event. More extensive information on the nature of seismic waves can be found in Richter (1958).
Rock acts as a low p~ss filter; low frequency waves pass with ,
little alteration, while high frequencies are attenuated. There is an approximate linear relationship between frequency and
attenuation. Energy attenuation is proportional to_the square of the distance traveled, pulse duration, and a number of environmental and rock properties. Environmental factors include pressure, temperature and saturation. Rock properties include density, seismic velocity and nature of discontinuities. Details of these factors can be found in Young (1986).
r Il
6.2 TD CAMPBELL DD LAD MID MONITOIUNG SYSDM
Campbell Red Lake has a 64 channel Electrolab MP-250 microseismic monitoring system installed at the mine. An original system used 32 veloci~y and 16 accelerometer transducers. The present system uses 64 accelerometers. Data manipulation and storage ia
11'0
,
o
c
performed with a POP 11/23+ computer. HareL copy is ~btained from a Teletype model 40 printer. Figure 6-2 i8 a schematic diagram of the system.
Transducer location depends upon monitoring priorities. Mine wide event detection as weIl ~ 'tight' monitoring of areas of
~~Jeater concern are possible. Source location accuracy improves r)1 ...
with the 'tightness' of the transducer array.
The MP-250 system sweeps the transducer input signaIs for signal voltages exceeding the preset threshold. The tbreshold is set to .. a level above mine cultural noise. Once the threshold ls exceeded, a time window opens and event processing commences. T4e time window can be set from 100 microseconds to 999 milliseconds. A-minimum of five transducer channels must be triggered within thë time window in order for event processing to continue. The time of the event is set when the time window first opens. At the end of the time wi~dow, information for that event is no longer collected and the window is reset for the next event. If fewer than five geophones are triggered, no event is registered. , The MP-250 is also equipped with an energy integrator. The signal from a preset"transducer passes through the integrator which calculates a relative energy for the event.
The processor receives the channel numbers and arrivaI times for each event. The transducer co-ordinates and the p wave velocity must already he entered into the computer. From this information, the source location is calculated by the solution of linear equations. The mathematics of these calculations are given by Blake (1982) and Electrolab (1981).
The PDP 11/23+ receives the data from the MP-250, it in turn calculates source location based on a least squares solution. The POP 11/23+ also calculates a standard deviation error for the solution. Interactive programs allow removal of spurious triggered geophones and other data manipulation to enhance foource location accuracy. Accurate source location by these methods
111
o o • "" ID c: ~
0
G) 1
1'1)
junctron ~I amplifier 1. Itra nsd ucerl .. box
i: li!. UNDERGROUND EQIPMENT
. () .. O-
• 0 MP250 MICROSEISMIC PROCE$SOR
• .. 3 .. N ()
i: g -.. 0 ..
1 I~ energy -filter volt
L circuit sensor
event dota . r -.
prodessing -r-- tlmer
. ,
:a ca Ct) 'IC • -• data 3 processing, "
POP 11/23 data printer COMPUTER storage piotter
\
T
t
.. •
(
(
assumes straight ray propagation which is only the case in
homogeneous media. In a mine, the radiated rays are distorted by ope~ings, changes in rock type, fracture zones and stress
changes. High energy events, occurring close to transducers, can be located with reasonable accuracy.' Low energy events and event~ occurring at a great distance result in po or arrivaI time data tiue to energy attenuation and ray bending.
Cultural noise derived from drilling and electrical currents is a
major problem. Although costly, sorne cultural noise can be
removed by notch filters. "Electrically induced 60 Hz and its
harmonies are easily remo~d. Drilling noi5~ i5 too broad band . -for simple notch filtering and i5 generally eliminated by
increasing the trigger level" however, increàsing trigger levels
results in the 105S of sensitivity to low energy events.
6.3 MONITORING DESTRESS BLASTING
T~e remainder of this chapter deals with microseismic monitoring
for determining the effectiveness of destress blasting. ~
Microseismic monitoring to determine the location of highly
stressed ground that may require destressing is beyond the scope
of this report. Information on this topic and other usês of
microseismic monitoring are available from Leighton (1982, 1984),
Ob~rt and Duvall (1957).
There are three possible mïcroseismic methods for tne analysis of
the effectiveness of a destress blast. The first is the
monitoring of the noise and energy emissions following the blast ~
and comparing them to the emissions before the blast. The second
-- is to conduct ~ seismic velocity survey be~ore and after the
blast. The third is to evaluate the energy released during and
. immediately fOllowing the blast, and equating this to a reduction
of the stored strain and potential energy of the rock.
113
J Source location is of importance to any monitoring program.
Before any data c~ be analyzed, its location must be determined. A tight array of geophones should be placed around any stope to be destressed to ensure accurate location of aIl events.
Lahoratory tests have shown that as the stress level on a rock increases, the microseismic noise rate increases. AS rock known
to have failed adjusts to a stable condition, the noise rate decreases. Laboratory results of Obert and Duvall (1945) are shown in Figure 6-3. These results have also been observed in the field, Obert and Duvall (1957). Experience at Campbell Red
Lake Mine has shown that areas can be stressed to the point of bursting with little precursory or postburst aqÙivity. In only
one case has a correlation between microseism~~ activity and instability been made, Neuamann and Makuch (1984). It is this authors opinion that with highly sensitive monitoring equipment placed in strategie locations, evaluation of stability can be
provided by microseismic monitoring.
The pre-blast activity level is usually higher than the mine
background due to the ground being overstressed. High microseismic noise and energy release rates are "expected to
follow a destress blast~ These are due to readjustment of. tb~
fractured ground to the stresses imposed upon it. The activity should diminish within a short time if the destressing has been
successful. If the microseismic activity remains high following
a destress blast, then the blast has had no effect or has reduced
the stability of the area. Destressing causes stresses to be
transferred to adjacent ground. The potential exists to destress
one area and create an overstressed area sorne where else. Stress transfer is best predicted by numerical modeling methods.
Rock noise rate is the most widely used form of analysis. The method is simple and requires a minimum amount of equipment for
collecting the data and for data analysis. Both single and
mult1~channel monitoring equipment can be used. The rate can be
analyzed in terms of number of microseismic emissions per u~it
114
C 60 ....
50 GRANITE ·A"
40'
• CI.I JO +' ~ c::
'r- 20 e s... CI.I 0..
'" 10
+' c:: cu > CI.I ,. 0 10 20 30 40 50 60 70 80, 90 100 CI.I
+'
"' 50 s... u
'r-- -e 40 GRANITE -B H
'" 'r-e cu VI 0 s-ur 30 .~
::E:
20
10
o 10 26 30 40 50 60 70 80 90 100
Applied load/ultimate strength. percent
Figure 8-3: Selamle Actlvlty Veraua load
~fter Obert and Duval 1946
c .~-
115
o
time, accumulated emissions and time betweenflmissions. It should be noted that an emi~sion registered by the monitoring system is dependent upon certain parameters of the particular system; these include: the threshold trigger level, the-frequency response, and in the case of multi-channel systems, the time window and number of transducers requirea to register an event.
Energy can be analyzed for release rate per unit time, accumulated energy and time between events with a specified energy value. Any analysis of energy must proceed with ca~tion, as_J:he computer calculated energies are relative values for the foilowing reasons; . "
i) Calculations based on a single axis geophone are unrealistic as seismic waves, hence energy, are three dimensional ..
fJ(
ii) Attenuation is neglected or assumed equal in aIl directions: this is seldom true.
iii) The time window may cut off part of the wave to the energy transducer, giving a reduced value.
Energy can still be used for analyzing destresB b~astsl ,although its limitations/must be considered. It is still possible to compare the relative energy of mlcroseismic events occurring in ~ose proximity to each other. The to~al energy of a destress blast can be estimated and compared to the theoretical energy re1ease of the explosives a10ne. Any energy above that of the explosives is energy--released by the rock mass.
Seismic velocity or wave propagation velocity.can be used to determine the stress conditions in a rock masSe The velocity at which seismic energy traveis through ~ medium iB proportional to the modulus of deforrnation. A reduction in the seismic velocity indicates a reduction in the modulus, hence, a reduction in stress concentration. Figure 6-4 shows a sei smic survey layout
- and results of moni~oring a destress bIaBt.
116
c
•
c
• i 0: ... ~
SECTION" A'
, ,
•
M .,A
NO, @t.
• G.o ......... 5700 levil
~ • el: • •
~ë~:··oœ~ 0'\.00
0
V.loclfy lU' "., fftol hole o o o o o o
• ; l' U
~ SHORE OESTRESSING
" IV ... , . • • •
o 10 ZO )0 . , , . AnER OESTRESSING
aelamle contour plota ln "feet per aecond
Figure 6-4: Selsmlc Survey layout And Résults
afte' Leighton 1082
..
117
•
)
•
6.4 SOJeGUa AND CONCLUSIONS 01' HIcaoSKISMIC MONITOlUNQ
The microseismic monitoring technique is an extremely useful tool fo~ ground control purposes, although like ~osy monit?ring equipment, its limitations must pe recognized. The problems associated with attenuation, ray bending, cultural noise and energy calculation need to be further addressed.
Microseismic monitoring can be used to asses the effectiveness of destress blasting in a number'of ways: monitor the stability of
1
the g~~und before and after destressing, monitor the energy . .
released during the destress blast and immediately, following it, asses seismic- velocity before and after the blast.
t:.
\
. -
. - ---~
118
( ~
c
(
7 rIBLD STOnIES or OESTRESSING HARROW VElM STOPE rACES
Many mines have utilized blast fracturing methods to control
high stresses at stope faces. The technique is to drill and blast short length holes in the breast face or out into the wall
• rock. The holes generally exceed the length of the production round by 1 to 1.8 m. Blasting is conducted either pre or
synchronous with the production blast.
In general, the practice has not been weIl documented, with the
exception of the early South African efforts, Hill et al. (1957).
In Canada, Kerr Addison used breast face destressing extensively
during square set mining, see Chapter 4. It is also known that
Dickenson, Macassa ~d Lac Shore mines used the method, although
documentation is min~mal. , Campbell Red Lake Mine (CRLM) has destressed horizontal breast faces in cut and fill stopes for a
number of years, ~etails are given in Chapter 4. The mine
recognized the need to gain a better understanding of the destressing process in order to evaluate its effectiveness and to
aid in optimizing blast design.
This chapter reports on two investigations carried out by the
author at the CRLM, mine site. The first investigation was
designed to evaluate the fracture zone created around a single
destress hole in a breast face. The second investigation
evaluated the influence of destress blast hole position and
overall stope reaction to destressing. The effectiveness of the
destress technique is also appraised.
7.1 rRACTURATION AROUNO A SINGLE DESTRKSS BOLI
The object of this study was to investigate the extent of
fracturing around a single confined hole blasted in a breast
face. A single blasted hole has no possibility of constructive
stress interaction with adjacent blasted holes. Other
researchers have undertaken"similar studies, as discussed in
119
l
)
o
~
Chapter 3, however, they have been conduct~in reasonably
isotropic rock with low in situ stress levels. The work /
conducted in this study comprised of observation and measurement
of fractures about blasted holes.
The stope selected for the study was 1702 lE 'B', located in the
'A'- ore zone on the 17 level. This stope had a history of significant sei smic activ~ty and was being destressed on a
re~ular basis' at the time of the study. Destressing was
conducted by blasting a 32 mm diameter, 4.5 m long destress hole
in the center of a 3 m long breast round. The destress hole was
fired with t~e round. After blasting the bottom 1.5 m of th~ destress hole could be examined. Only six destress holes were examined as mine scheduling caused tne project to be
discontinued.
After blasting, the destress hole diameter was observed to
fluctuate greatly. In one case the hole was completely closed
off; the driller reported'very easy drilling in this face. The
maximum diameter at a depth of 0.45 m from the new face was 43.2
mm, the average was 38 mm. These measurements were taken in. the
direction of maximum elongation. At the collar the hole wouid
usually be belled. Figures 7-1 and 7-2 show the variation in • direction of elongation. In Figure 7-1 elongation i5 in a
direction at high angle to the vein structure, but still paraI leI
to a pre-existing discontinuity direction. In Figure 7-2
elongation is parallel to the direction of the vein st~ucture . •
Figures 7-1 and 7-2 also indicate the discontinuous nature of the
rock in the study area. Major discontinuities are perpindicular
and paraI leI to the strike and dip of the ore body.
A crushed zone was observed to extend from 3 to 10 mm around the
destress hole periphery. In cases where the hole was belled and
could not be scaled it was not possib~.to get accurate measurernents. Brady and Brown (1985)' suggests that the crushed
zones should extend out to one bore hole diameter, in the case of
,~CRLM, the strength of the rock appears to limit damage.
120
• Figure 7-1: Blast Hole Elongation in Horizontal Direction
c Figure 7-2: Blast Hole Elongation ParaI leI to Ore Vein
c 121
...
)
.Visible fractures were non-existent in sorne instances but generally extended from 50 to 100 mm (1.8 tOI 3 borehole diameters), as shown in Figures 7-1 and 7-2. The fractures were of varying orientation but always followed along pre-existinq discontinuies.
....1
In order to gain a more accurate measure of the cru shed zone and extent of fracturing, attempts were made to force diluted masonry paint into the fractures. The first method use~ was to fill the destress hole with paint then blast with PowerrneK cartridge explosive. This proved to be unsatisfactory as the paint appeared to becorne involved in the explosive reaction. The second method was to fill the destress hole with paint after being blasted but prior to blasting the breast round. A pluq and header, similar to those used to plug diamond drill holes, was installed at the collar of the hole, and the hole w~s pressurized to approximately 600 kpa with compressed air. Good results were
obtained with this method although after blasting the round, the ' paint tended to spill out making ~racture identification in the
lower portion of the hole difficult, Figure 7-3. A crushed zone clearly extended out to 10 mm and a fractured zone to 50 mm in a number of radiating directions. At this point the project was terminated. The next blast was to use red paint, instead of white; as weIl, after pressurizing ex cess paint was to be flushed from the hole.
General comments solicited from the driller were that he liked the ide~ that destressing may prevent rockbursts. He reported that destressing had no influence on subsequent drilling.
7.2 zrrZCTS or DZSTRBSSING ON STOPZ REACTIONS
The second single hole destress blast monitoring program was conducted in 1704~W sto~e, located in the 'Al' ore zone on the 17 level. The program was designed to ass~ss stope reactions, and the effectiveness of breast face destress blasting. In
122 ..
Figure 7-3: Blast Hole After paint Injection Trial
•
c i r
. '"
o
c 123-
'. ;.
o
o
addition, the differences between destressing a back hole or
center hole were ~luated.
The 1704 EW stope was selected for the study for a number of reasons; it was experiencing a large number of seismic events suggesting high stress, percent age extraction in the vicinity was high, the stope was being mined through a section with a
~
relativel~ consistent width and geology. The stope and the location of instrumentation and monitored breast faces is shown in Figure 7-4. Figure 7-48 is an enlargement of the test area. The notation on these Figures is N for no destress hole, B for back destress hole and C for center destress hole.
A total of 10 cQnsecutive breast blasts were taken during the study; 4 with no destress hales, 3 with the back hole as the destress hole, and 3 with a hole 0.6 to 0.8 m from the back as the destress hole. Face dimensions were: width between 1.5 and 2.4 m, height between 2.4 and 2.8 m. Breast length was 3 m, the dèstress hole was 4.5 m long. Production and destress holes were 32 mm diameter. The destress hole was incorporated into the breast round and blasted as such with AN/FO, detonated with safety fuse. A typical blast pattern is shown in Figure 7-5; this blast did not contain a ~estress hole. A destress blast pattern would have either the center of back hole extended.
Blast and stope monitoring inciuded a microseismic monitoring system, tape extensometer and visual inspections after each blàsts. Vibrating wire stress gauges, fill pressure cells and stope convergence met ers previously installed in the stope ~ere also monitored. Minera comments were solicited.
7.2.1 NICROSBISNIC MONITORING
The stope was instrumented with a tight array of transducers, the closest ones weQe located in 1604 W. drift and 1602 W. drift as indicated in Figure 7-5. Monitoring consisted of recording •
124 /
<,
z
C fi ::
GI 0 CI CD :J ..
! al ~ -- CI • ca 1! 10 • CD Q. 10 ..
G Q.
! ;
_ ........... _--11-1 ..
c
c 1
~ ~
FIgLn 1-4: 1104 EW Stope, Test Location And instrumentation ..
125
o ,
DO
r
o
Flgwe 7-48: Test Blut 8equance
o 1,26
t,
- -" T,
Figure 7-5: ~ypica~B1ast Pattern in Narrow Vein Stope
•
(
•
c 127
_. __ ._-----
..
\
everits occurring in the stope as well as keep±ng a record of the number of hits to each of the four closest transducers for three and one half hours after each blast. As discussed earlier, the object of destress blasting is to fracture the ground. This being the case, it is expected that the m:J,.croseism'ic system should record -0/ greater number of low energy events as slippage occurs along ffactures. High energy events, caused by the formation of fractures, are expected ta be less prevalent with ~he use of destress holes. Figure 7-6, presents a summary of the activity from the stope after each blast. Low energy events were those whose energy determined by the MP2St was less than 100,
high energy events greater than 100; the recorded energy values are relative magnitudes only. The nurnber of hits is the sum of the hits to aIl four transducers. \ .
The three types of seismié activity analysis (low energy, high
energy, total hits) aIl followed the sarne trends in terms of number of events. Monitoring errors causea by the following factors would introduce sorne error: triggering events by blasts
in adjacent stopes and rnasking events by blasting or seismic
events in other stopes. There was no distinction between the different blast designs based on anàlysis of seismic acti~ity.
In blasts Cl and N4 reblast, seismic events with MP2~0 energy greater than 500 were recarded. This i5 approximately equivalent to bla5ting 1 kg of AN/FO. Although the mlcro5eismic system dete_J:"rnined the è'vents to be at the breast face, in nei ther case was any visual evidence of the event detected .
7 .2 • 2 STOPB CONVBRGB!NCB •
Stope convergence monitoring was originally conducted witn RST Ltd. electrical convergence monitors. This equipment broke down and was replaced by a tape extensometer w1th eye-bolte installed 0.3 m into the wall. The RST equipment 18 more accurate,
however, it i5 not very rugged; it i8 designed for permanent
128 ,,'
·c
c
,-
110
100
go
80
~
~ 70
0
.2 60
e -i 50 • e u 40 Ë ~~
30
20
10
0
e:2I hlt8 (x10)
N1 S1 C1 N2 92 C2 . bla8t type
ISSI evont. (low)
N3
Figure 7-6: Mlcroselsmlc Actlvlty After Blasting
•
83 C3 N4 N4
l?2ZI eveti~ (hlgh) •
7~-------------------------------------------------------'
8'
1 ~ " E S -
1 1 •
O~------~~~----~~------~~~~--~------~~------~
~1 + pt. 2
Figure 7-7: Stope Closure
12"9
B3 C3 N4
l pt. 4- x pt. 5
)
o
installation such as in backfill. The accuracy of the tape ext~nsometer is 0.5 rom.
The location of the monito;ing points is shown in Figure 7-4b. /
--In aIL cases stope clos~re occurred after blasting. Closure wAs recorded until the breast face was 3 to 6 m ahead of the
• monitoring p0~nt. It was expected t~t the initial closure would be greatest for destressed faces if they were successful in creating an increased fracture zone. Because of the failure of the electrical convergence rr.eter, data is only available for 6 of the 10 blasts. Figure 7-7 is a graph of the measured results.
~ Instrument installation and measurement was conducted as soon 'after the blast as possible, however, it is likely that the greatèst closure occurred prior to installation. The high convergence at point 4 is attributed to the eye bolts not being anchored in solid rock. No distinction can be made between destressed and non-destressed blasts on stope convergence.
7.2.3 VISOAL INSPECTIONS
After each blast 'the stope was visited; records were made of the condition of the blasted area including such items as stress fracturing, the amount of sca1ing, seismic activity, bootiegs, condition of destress hole, miners comments and any other
·anoma1ous conditions. It was expected that destressing wou1d decrease the amount of stress fracturing and a1so reduce spitting from the face. Based on the information c01lected there i8 no perceivable difference in the different b1ast designs.
Stress fracturing extended up the face, from the brow, 0.9 to 1.7 m before scaling and 0.6 to 1.2 m aftèr sca1ing, typica1 stress fracturing is shown in Figures 7-S"and 7-9. The fa ct that much of the stress fracturing is superficial and easily 8caled' away suggests that it May be blast induced fracturing.
130
•
«
-Figure 7-8: Stress Fractures Before Scaling
p~.r , ,f." "l
,j .. J.
"
Fractures extend to within 1.5 m of the back . .,;1
Figure 7-9: Stress Fractures After Scaling Fractures restricted to vicinity of brow.
131
)
-
o
The hanging wall was defined by joint contacts running subparallel to the vein. For blasts NI to B2 foot wall contact was a~ong a shear plane, there was no pronounced foot wall
for blast C2. Foijblasts N3 to N4 the foot wall broke jo,int contacts similar to those on the hanging wall.
contact 1 along
After blasting the destress holes wertenlarged and generally
belled at the mouth. A crushed zone extended from 8 to 29 mm (rneasured after scaling). In sorne cases the h~le was ovoid with the long axis vertical or horizontal. Both vertical and
horizontal open radial cracks were observed, these were always
related to pre-existing discontinuities; The B3 destress hole ls
shown4ln Figure 7-10, a vertical crack runs parallel to a quartz-, calcite infilled joint. A horizontal fracture is shown in Figure
7-11, blast B2. Closed radial fractures not associated
with known discontinuities, extended from 50 to 150 mm, fractures
associated with pre-existing diséontinuities extended much
further. Fractures were best observed after being sprayed with paint.
Much of the information proved to be ambiguous and very difficult
to interpret. For exarnple, it was found that in blast Cl there
were ;ery few stress fractures and a large amount of seismic
activity during the shifti after blast N3 there was little stress
fracturing and only a small amount of sei smic activity during
shift. The greatest amount of stress fra~turing and on shift
seismic activity occurred after blast C2, this was located
adjacent to the highly jointed and fractured pillar separating
the off shoot ore vein. Microseisrnic activity recorded
irnrnediately after this blast was not high.
Ovoid holes, along with the greatest crushed zone around blast
holes, occurred after blast B3 and C3. These indicate the
presence of high stresses. In the case of B3, extensive
microseisrnic activity was recorded after blasting, much less was
recorded after the C3 blast. The on shift seismic activity was
similar in both cases. ,
132
c Figure 7-10: Blast B3 Destress Holé
Open fracture runs parallel to ore ~~in.
Figure 7-11: Blast B2 Back perimeter Hole Fracture runs sub perpendicular to vein direction.
c
, - 133
•
,
The drillers reported having difficulty drilling holes in
fractured ground below the previously blasted destr,s.ss hole.
Holes drilled immédiately below the destressed hole were the
most difficult. This observation contradicts that of the 1702 E stope study. Two possible explanations are: differences in rock
ty \or differences in miners sensitivit8 to such factors. Other
wo has indicated that fracturing in Campbell Red Lake Mine
extends to 15 blast hole diameters, Chapter 8. 1 -
l f~actures can also be expected around the brow where
the compressive waves are reflected as tensile waves .
. Damage to the stope back due to back destress holes was noted in
only one case. Addi~ion~l scaling was r~quired which ~ft the back with an unfavorable slot shape. After a few days and more
scaling, the slot was no longer prese~t. After project
completion back degradation was inv,estigated. The amount of
100se in the screen, and general condition of the rock was noted.
Breasts blasted with center destress holes had the least amount
of back degradation. Breasts with no destress hole or back
destress hole generally had 5 lo 15 cm of loose in the screen.
It must be recognized that this study is subject to the quality
of the original scaling and whether or not lObse was removed from
behind the sereen at a later time.
7.2.4 OTHER INSTRUMENTATION
Vibrating wire stress rneters, backfill pressure cells and
convergence meters installed in backfill were monitored during
the test. The position of the instrumentation is shown in
Figure 7-4. The instrumentation was ~ocated at distances too
great to diseern the different blast types.
As mining progressed under the stress g~uges no change was
recorded in gauge l, a stress reduction ~as recorded in gauge 2.
Gauge 2 was approximately 6 m above the mined lift. It was
expected that an increase in stress would be registered as the
134
(
(
c
distance is beyond that normally associated with excavation damage and stress relaxation. It is concluded that the wedge holding the gauge in place was loosened by blast, vibrations.
The fill pressure cell and convergence meter registered no change
after being mined over. They were located approximately 25 feet
below the lift back. It is thought that initial closure occurs when ground is first opened, however, no further closure occurs
until after a much greater area of ground is opened.
-7.2.5 DISCONTINUITY SORVEY
t In order to assess the influence of geological discontinuities on
the microseismic activity, the study area was mapped for joi~s, fractures and other pertinent information. AlI together 7
discontinuity orientations were identified, listed in table 7-1. l
A substantially greater number of low energy events occurred
a~ter passing through a zone of high joint frequency. The high
discontinuity zone acts like a destressed area, non violent
yielding may be occurring, resulting'in an increase in low energy
seismic emissions. Discontinuity orientation did not appear to
influence microseismic activity.
Figure 7-12 presents the results of the mapping. The respective
discontinuity traces are marked on the hanging wall, foot wall
and back. Stress fractures in the back generally broke along the
major joint breaks, these have not been marked on the figure. It
was common to find large differences in the hanging wall and foot
wall discontinuity patterns. In the area of blast Cl very few
joints penetrated the foot wall shear conta the area
of blast C3 the foot wall rock was extensively fr tured. This
is a result of it being in the pillar of the off shoot ore vein.
135
o
\
o
o
Tabla 7-1: Dlacontlnuity Orientation
Type 1: strike 260 deg. dip 30 deg. N., joint set quartz-calcite infilled a - 1 mm
Type 2: strike 280 - 315 deg.,dip 80 - 85 deq. NE. joint set, quartz-calcite infilled 0 - 20 mm
'shear contact ore vein cleavage.
Type 3: strike 60 deg., dip 85 deg. SE. joint set
Type 4: strike 350 - 005 deg., dip 20 E. joint set, trace quartz-calcite infilling
Type 5: strike 350 - 000 deg., dip 85 deg. E. joint set, trace quartz-calcite infilling
Type 6: strike 180 deg., dip 30 W. joint set
Type 7: strike' 70 - 80 deg., dip 30 deg. S. joint set
(
136
( hanging wal foot wal back
(
1 1
1
1 1
Il ~ 1
1
1 1
c FlgLra7-12: Dtacontlnulty Map
~ 37
,
l
o
7.3 SUMMARY AND CONCLUSIONS OW BRBAS~ ~ACB DISTRKSSIHQ STODI.S
Single holes blasted as destress holes in narrow vein stopes resulted in a number of observations of blast induced damage and response of the stope to destress blasting. From these observations general conc~sions on the effectiveness of the destress blasting technique can be made.
The studies indicated that a blast induced crushed zone extends 2 to 30 mm about the blast hole. The crushed zone was circular and apparent1y not affected by pre-existing discon~inuities or stress regime. The variation in extent of the crushed; zone is attributed to the strength of the rock about the blasted hole.
A blast induced fracture zone extends from 0 to over 100 mm about ~the blast hole. The fractures a1ign themselves along preexisting discontinuity planes. The influence of in situ stresses could not be determined. Blasted holes tended to be ovoid with the long ~xis parallel to the orientation of maximum fracture extension. Blast induced damage may extend beyond the visible fracture zone as has been obser~ed in other studies, Chapter 8. Damage is dependent upon
4
the proxirnity of a free face, the preexisting fracture density, in situ stresses and the strength of pre-existing discontinuities. A destress hole placed towards the
r'\ back will not induce fractures as a\result of tensile reflection of the stress waves at a free face. This view is based on theoretical considerations of stress wave attenuatio~tlined in Chapter 3. The creation of an adequate sized cushion zone la likewise not possible.
The study revealed no significant differences, relating to a reduction in rockburst potential, between breast blasta with a destress hole and those without. It is possible that differences did exist but were beyond the range of detection of the
instrumentation used. Attempts were made to reduce the number of variables duting the tests, however, variables su ch as geology and in situ stresses could not be controlled. The number of pre-
138
c
(
(
..
existing discontinuities was observed to significantly effect the microseismic- emission rate after blasting.
The normal'extent of damage;about a stope May well extend beyond the distance of any additional, destress blast, induced damage. This bein? the case, it is unlikely that a 4.5 m destres8 hole in a 3 m breast would have any significant effect on stope reactions.) This May expl~in why damage i8 seldom seen in a
/ breast J8ce after high energy sei smic events; the events occurred at a distance beyond the existing fracture zone.
Future work should investigate the extent of the blast induced fracture zone about the breast face, if this distance exceeds 1.5 m destress holes should be drilled to fracture the solid rock beyond. The location of high stress areas should be determined for hole positioning, this could be done with the microseismiç system by monitoring the exact location of seismlc energy being released.
Based on the information collected during thls study, blasting a . . 4.5,m destress hole in a 3 m production round is not an effective
" means of destressing. This conclusion i5 specifie to the test. environment; changes in geology or stress regime May yield different result5. It ia believed that a single 32 mm diameter hole in a 3 by 2 m face does not impart enough energy to the
groupd to cause adequate fracturing. In order, to be effective, the ground must be hit with sufficient energy to cause extensive fracturing, not just local. Fraèturing and reduction in modulus must be created throughout the entire breast face'area.
• J 139
1
o
8 BUST l'RAC'l'URING STODl:ZS r
During'the course of research at Campbell Red Lake Mine (CRLM) , experimental studies on blasting and its influence on the rock ' mass were carried out at the mine site. The experimental studies were designed~to investigate a number of factors of importance
- " to destress blast design; the attenuation of blast induced seismic waves was investigated in an attempt to estimate blast induced fracturation; the extent and nature of fracturation
between two blasted holes was investigated to determine actual fracture extension and reduction in modulus of deformation caused
-by fracturation. One of the planned studies involved an 'intensive investigation of a stope pillar before and after destressing. Because of a change in mine scheduling this study
, could not be completed. '\
, 8.1 l'RACTURATION STUDIES IN 'l'BE 1670 W 'A' S'l'OPB CROWN PILLAR
The 1670 W 'A' stQpe crown pillar is located in a miscellaneous mining zone on the 15 level, Figure 8-1. Original plans for this pillar called for destress blasting followed by underhand cut and fill mining. The pillar had not shown any signs of being over stressed, however, production supervisors wished to destress the
pillar as a safety precaution. As a result of a change in mi~e scheduling the planned study for this pillar could not be completed.' One hole was drilled thro~gh the pillar resulting in . . information on the relationship between ex~avation induced fracturation and modulus of deformation.
The anticipated destress blast design, for the pillar, was similar to that used in the 1604 EW and 1802 E destress blasts, Chapter 4. Site investigation plans included drilling 76 mm boreholes through the pillar before and after the destress blast. Probex-l dilatometer surveys would be run in these holes'to obtain modulus of deformation and fracturation data. The change in mine scheduling resulted in only the pre-destres8 blast hale
,:-J 140
\
f . \ .~
, .
(
, . \
. ,
il.
, ..
1" 1 & Lev.1 Plan FIGur. 8-" .
f
_3.é::.17.LN3/1 J.N!Or ~.
4"
141
)
Q
•
l
-0
being drilled and logged. The location of the hole, and pillar geometry a~e shown i9 Figure 8-2.
The results of the dilatometer survey and core analysis are presented in Table 8-1. As was expected, the c~nter of the pillar contained few fractures. A greater nurnber of fractures were detected at the two collars .. These fractures were the rësult of excavation damage and high stress. Reasonable correlation exists between decreasing modulus and increasing fracture density as indicated in Table 8-1. Dilatometer results around the collar cannot be substantiated as the instrument was cauaing rock movement along discontinuities. This was detected
~ b~ audible 'snaps' and, in one case, observation of moveme~t.
~The core is shown in Figure 8-3. Selected core samples were tested for uniaxial compressive strength and Young's modulus, Table 8-1. It is dirficult to' compare dilatometer modulus of
deformation and laboratory derived modulus for the following
reasonSi the two methods measure deforma~ion in different directions, laboratory samples are intact while the dilatometer measures over, and includea in, any reductions due to fractures; laboratory moduli determined in the laboratory w~th strain gauges are unlikely to include effects of fracturation due to the small length over which they measure. ~
Two horizontal and two vertical strain gauges were mounted on
each core sample tested. Poisson/s ratio was determined to vary
from 0.19 to 0.25, the mean being 0.21. It was observed that Poiason's ratio increased alightly with confining pressure. This is expected when loading is para11el to major discontinuities, as
was the case. A s1ight increase in Poisson's ratio a1so occurred ,
with increasing fracture density. A typical stress strain curve and horizontal versus-~ertical strain curve is shown in Fi9ure~
8-4 and 8-5.
Flgur • • -~: 1670 W"A' Stop •
•• ctlon vl.w
c
• .. - 0 = .. .... ... (1)
~ -- , ~ 0 .... .. eo 0 .... • c
0 .. N • = -. \ IL
c: • 0 .. 0
C ,-
l 0 ., (1) , 4(
~ 0 r-. G ..
= -...
. .
,
1
\ )
}
Table 8-1: 1670 W'A' Stope Crown Pillar Test Results
~ • " dist., fracture (~-1)
modulus (GPa) strength Poisson's from density u.c.s. ratio collar dilate core dilate core (MPa) 50% U.C.s. {~m}
• 139 16 2 ./
190 10 Il 260 4 75 134 0'.22 266 5 ,25 295 2 84 199 0.19 380 7 85 123 0.12
• 505 3 86 ' 176 0.21 520 3 97 647 0 94 901 3 -68 915 3 87 148 0.18 978 7 37 1043 7 49 1060 10 88
'( 166 0.24
:> 3 41 10 10 14 10 10 10
5 8 87 188 0.25
•
o
144
.'
c
Figure 8-3: Core Recovered From 1670 WA Stope Crown Pillar Extensive fracturing occurred at both collars •
• l
•
.. 145
•
~
-1It . .iJ
··0 ..
.~
FIGur. 8-4: Str ••• Straln Curv. CRLM And •• It •
170
160
150
HO 130
120
• lUI A. :1 100
• 90 • • 80 .. .. 70 • 60
50
~,
, JO
20
10
0
0 00002 0.0004 0.0006 00008 0.001 0.0012 00014 0.001 li 0.001'11
Figure 8-5: Horizontal ver,uI Vertical Str.ln
II: -• .. .. • • .. II: o ~ .. o ~
0.0005 ..,-------------------------"':1..
00004
o.ooo~
00002
0.0001
0 0 0.0002 0.0004 0.000& 0.00011 0.001 0.0012 0.0014 0.001' 0.00111
.ertlca' .traln ..
/'"
1'6
\.
c
c
c
Compressive failure of most samples was along pre-existirig
discontinuities, 0 to 10 degrees to the core axis. A few samples failed along planes at 10 to 25 d7grees to the core axis; this
failure was of ar_~w tooth nature across the pre-existing 1
discontinuities. Two samples failed violently, although there
was no apparent feature that distinguished these samples from the others.
The central portion of the pillar was relatively intact. Few
fractures exist beyond the pillar boundaries. Analysis shows a
fractured zone to exist to a depth of 2 m from the top of the
pillar and 2.5 m from the bottom, Figure 8-2. The fractures are attributed to excavation damage or tension fractures resulting
from high horizontal stresses. Stress relaxation may be the
cause of sorne fracturing in the immediate vicinity of the opening. The greatest fracture frequency and least RQD occurred
within the first meter away from the openings. Fracture densities
greater than 15 and rock quality designation (RQD) of 0 percent
occurred in these regions. A transition zone of increasing RQD ~
and decreasing fracture density lies between the ~ighly fractured
zone and pillar core. Fracture density in the central region was
from 0 to 7 based on core analysis, RQD was 100 percent. Only
one bore hole impression was made in the pillar center, no
fractures were detected. Although a double barrel core recovery
system was used it ls possible that m?ny of the fractures
recorded from the core were drill !nduced. Nin~ty percent of the
fract~res recorded by core analysis and impression packing ~
occurred along known discontinuity directions. Many of the core
fractures were quartz carbonate or phyll~ilicate infilled.'
8.2 rRACTURATION STUDIIS IN THE 1470 STOPE ACCESS DRIrT
The 1470 Stope access drift is located 645 m below surface on the , 14 level, Figure 2-3. The study evaluated blast fragmentation
between adjacent blast holes.
147
,.. The 1470 drift was selected because of its ease of access, proximity to an in 9itu stress measurement location, and remoteness from mining activity. Severe rockbursting has
occurred 150 rn west of the test location, Hedley et al. (1985), suggesting the area is heavily stressed. Virgin ground stress
rneasurements conducted by CANMET, Arjang (1986), were taken 400 m south of the test site, on the sarne level. A 23 MPa horizontal stress was rneasured in a northeast -'southwest direction, the
other two principal stresses were 16 MPa, refer to Chapter 2.
8.2.1 ROCK CHARACTERlZATION
An extensive investigation of the rock properties at the test
site was undertaken. The rock was c1assified by the mine as an altered andesite. Typical modal composition is as fol1ows: biotite 5 to 15 percent, calcite 20 to 45 percent, chlorite 25 to
30 percent, feldspars 8 to 12 percent, muscovite 0 to 4 percent,
opaques 2 to 6 percent, quartz 12 to 18 percent, and fine matrix
,
15 to 50 percent. The rock was subdivided into units A to F ~, based on modal composition, degree of foliation, and number of cross cutting joints. The units were sub-parallel to the major
cleavage direction and andesite - altered rock contact. Detailed
thin section descriptions are given in Appendix 1. For
simplicity, the mechanical properties of the different units were
assurned the same. Subsequent laboratory testing proved this to
be a valid assumption.
The rnechanical properties of rocks are dependent upon a number
of variables. These varIables include mineralogical composition,
structural and textural features, crystal size and orientation,
stress level, temperature, degree of saturation and testing
method. Significant variation of these factors' can occur
over a very small distance. Assigning absolute values of rock
properties is therefore not recommended. Values should be given
a range based on statistical analysis of collected data.
1..J8
( ,
c
Table 8-2 list the rock properties at the 1470' drift test site.
The methods used to obtain the results are also listed. AlI the testing methods assume a homogeneous linear elastic and isotropie
rock maSSe
In the case of blasting, dynamic stresses are produced by the initial detonation pressure while a quasi-static pressure is
generated by the ensuing gas pressure. The way in which a rock
reacts to dynamic stresses or static stresses ls different. When
considering the dynamic str~ss, dynamic elastic constants govern
J Ck behavior. Static elastic constants govern rock behavior
der the quasi-static gas pressure.
Dynamic elastic constants can be determined by either resonance
or pulse methods. The resonance method involves exciting a
specimen to its natural frequency then measuring wave length and frequency. The pulse or ultra-sonie method requires measuring
wave velocity and material density. Static elastic constants are
determined by loading the rock and monitoring deformation.
The method of determinfng the two types of elastic constants
differ considerably. Static constants are determined with slow
loading rates and high stress and strain levels. Dynamic
constants are obtained at high loading rates but low stress and strain levels. \ A comparlson of values determined by the the two
methods is only viable if done at equal stress levels. In
homogeneous, isotropic rock, the methods should yield similar
results.
An extensive literature review conducted by Lama and Vutukuri
(1978), found static constants to be from 0 to 30 percent less
than dynamic constants. The difference is attributed to the
greater influence of discontinuities on static constants. Highly
compacted rocks had similar dynamic and elastic values.
Different elastic constants are effected to differènt degrees by
parameters such as discontinuity density, saturation, stress
levei and testing direction. Kaneshiro et al. (1987) detèrmined
149
1
"-i
• Table 8-2: 1470 Drift Rock propfrues #
parame ter m~thod value no. samples standard (me~n) tested deviation
Young's static 84 GPa 10 . 3.9 GPa tangent (47 mm core) modulus
Dilatometer Probex-l 62 GPa 11 22 GPa , deformation (borehole ~ modulus dilatometer)
Young's dynamic 96 GPa * modulus (ultrasonic)
o~ poisson's static 4 0.03 . " (47 mm core) ratlo
Poisson's dynamic' 0.22 * ratio (ultrasonic)
density, 2.83 g/cc -~ unconfined static 151 MPa 15 28 MPa
compressive (47 mm'and strength 26 mm core)
tensile static 15.6 MPa 5 strength
,
* Moh~nty 1988
• /
150
•
c
c
~ . that the in situ seismic modulus was less than the laboratory derived static modulus, but greater than the in situ dilatométer modulus.
8.2.2 BLAST rRACTURATION
As was discussed &t length in Chapter 3, rock response to blasting is not well understood. Explosive interaction with the
rock mass has yet to be fully characterized, as such, accurate prediction of rock mass response to blasting by numerical
modeling is not yet possible. Examination and measurement of in
situ blasts can lead to characterization of rock response to
blasting, which in turn, can result in empirica~ predictive capabilities. A predictive model developed in this way is very site specifie due to the variable nature of different rock
masses.
Experimental test blasts were conducted in the 1470 dr~t wall.
Horizontal, 45 mm diameter holes were drilled perpendicula~ to
the drift wall to a depth of 2.5 m, Figure 8-6. The holes were spaced at 0.3, 0.6, 0.9 and 1.2 m. The bottom 1.2 m of each hole
was loaded pneumatically with Amex blasting agent (AN/FO) and
initiated with Nonel ms delays (numbers 0, 2, 4, 6, 8). prev!ous destress blasts at éRLM have utilized Amex in 45 mm
holes. After blasting, NQ (76 mm) diamond drill holes were cored
midway between blast holes, Figure 8-6. Diamond drill holes 1 and 2 were stopped after 1.2 and 1.4 m because of drilling difficulties.
AIl diamond drill hole cores were logged for fractures, joints
and petrology, and tested for point load strength, sei smic
velocity, and uniaxial compressive strength. A RocTest Probex-l borehole dilatometer was used to measure in situ deformation
modulus and make bore hole impressions. A complete description
of the instrument and testing procedure is given in Chapter 5.
fractured sections repeated dilatome~r tests were not performed
151
Figur. 8-8: Bla.t Hole and Dlamond Drill Ho" Looatlon. 1470 Drift
o
•
o
4860N
0.& m
O. OESTRESS H~
1 • ODH
o \
152
(
C
(
Where 'fracturing and spalling was excessive, dilatometer testing wa not possible. A graphical summary of the test results for bore hol 3 and 4 is given in Figures 8-7 and 8-8.
The core RQD and percent recovery as well as fracture measurements correlate weIl with the location of the blasted zone. The point load strength does not exhibit any clear trend. Because of a limited number of suitable test specimens little can be said about the seismic veloC'i,ty. Fra,cture and joint counts were taken over 10 cm intervals then normalized to count per meter. Figure 8-9 shows the core recovered from hole 4. A large amount of fractures occurred at the collar; a result of both the initial excavation damage and this blast study. The core from the toe is intact, it appears that blast induced fracturation does not extend far beyond a line drawn perpendicular to the end of the explosive charge. A 15 cm section of core, indicated in .. Figure 8-9 appeared to be disced. This phenomena is usually
attributed to stress relief of the cored rock; discing is
considered to be indicative of high stresses. Considering the fractured condition of the surrounding rock it is unlikely that
high stresses are the cause.
Samples from the core were observed under the optical and electron scanning micr~scope. No microfractures or other stress related featur~s were observed. It is suggested that aIl the blast induced fractures in the 1470 drift test site were formed along pre-existing discontinuities. This is in agreement with results obtained in the narrow vein destress blast, Chapter 7,
b and the results of other researchers who report that existing fractures grow in preference to new ones and ~ause a fracture free zone in their vicinity.
Bore hole fracture impressions were measured for orientation persistence, and intensity. It was not possible to measure orientation of the short persistence fractures. Very short
_ fractures (less than 3 cm) grouped a10ng the top and bottom of the borehole; a satisfactory explanation for this has yet to be
153
... Ut
•
l: ... Go w ~
,.
o
F
E 1
[ 1-C'
A
ROCK TYPE
r
o 50 100
ROO
050100
RECOVERY - Ci 50 100
MOOULUS GPa
"'-
150
, 0/
l
l o S 10 15
la 47.5 M~
o b
o o
& 7
VELOCITY km/S
6
IMPRESSION DRILL CORE FRACTURES JOINTS
m-1 m- 1
,.
•
b '-~ , :j
~~ b :J
h JO' • o 5 10
DRILL CORE FRACTURES
m-1
..... ~
w Z o N
o UJ )-
'9 .- . - --. ID
...
~ G' 1: ~ • C» t ... .. ... .,. ... o o .. :: ".
Q • • :: • • • :r Il :ID • • 1: ~ • Z o i c.t
~I i
\
.. Ut Ut
:s: 1-0.. W 0
~
......
IF
1 E
. 0
'e
B
lA1
lA
" ROC'K TYPE
\ --.J
f.
,
-;
. / 'J ,
~ 1-
f\
)
L cl .
l
-
} 1 - -
•
~-\ - ~
o 50 100 0 50 100 0" 50 100 150
Rao
,
~RECOVERY 010
MODULUS GPa
~
'\
i J
1
~
Is 47.5 MPa
l:
o o Cl
o
o
o o 87
1 a-L-o 5 10
VELOCITY IMPRESSION DRILL CORE km/s FRA~~r~S JOIN~~,
DRILL CORE FRACTURES
m- 1
~ -
-n G c: .. . '
• • • .. .. ,. .... 0
0 .. = 0 • • .. .. • • • -::1 CI
w :II z • 0 • N c: -Q .. • W •
1- :z: Cf)
-< 0 ...J -al •
~
•
•
Figure 8-9: .1
Core Recovered From 1470 Drift, Diamond Drill Hole 4 The core is disced over a 10 cm interval at the start of the third run. Dises were between .4 an 1.2 em thîck.
--
156
(
c
/ _/
determined. Of the measurable fractures, 80 percent were aligned . with known discontinuities, the remaining 20 percent appeared random. however, it was noted that most aIIigned with other pre-
e -
existing disco,ntinuities observed on the core. Stereographic
plots of the fractures measu~ed from the bore hole impressions and from tee core are shown in Figures 8-10A and a-lOB. The bore hole impression maps are located in Appendix 2.,
~t was observed that fracture frequency determined from the core
waS 30 to 60 pe~c~nt less than that recorded by the impresston
technrque. A plausible explanation for this is that the high
pressure u~d to create the impressions opens weaker fractures; drilling induced fractures may also be occurring. Many of the
drill, core fractures could not be identified' on the impression
map~, presum~bly these were due to drilling induced breakage (a single core barrel was used for recQvery). Attempts were made to
-
complete fracture traces to aid orientation procedures and
eliminate the possibility of counti~ a crack more than once.
Fractures less than 25 mm were very difficult to deal with.
Fractures parallel to the drill hole axis may have introduced
sorne bias to the results, however, very few of these were
observed in either the core or on the impression maps.
Figure 8-11 shows the relationship between fracture frequency,
from the impressions maps, and distance between blast holes.
Figure 8-12 shows the relationship between dilatometer modulus of
.deformat~on and distance between blast holes. Distance between
b1ast holes is measOreçi a.long a 1ine perpendicular to the diamond
drtll l~ole. No adjustments have been made for the diarnond drill
holes not being exactly c~ntered between the blast holes. The
9~neral trends are decreasing fracturation and increasing modulus with increasing distance.
Figure 8-13 demonstrates the general reduction of in situ
deformation modulus with increasing fracturation. Current CRLM
practice has utilized 1.3 to 1.5 m destress blast hole spacing.
Based on the experim~tal results, at 1.5 m spacing no
157
• Figur. 8-10: 1470 Drift Fractur ••• Imp ..... lon Pack.r
\
o 158
-Ut ~
---
II'
+
• ~I
5-9 À-Z
~
\
1\.
N ~- r
/
Î/·/'
++ l, + +
s
} r
"
'+
~
" :
"II II c ; • • ,-0 Il •• -... -III 0 G .. -... 1 • .. .,. ; ft
~ ... . , • 1
"0 :: II: n 0 .. •
~
.r
•
0
..
Figure 8-11: Fractur •• ver.ue BI •• t Hole Spaclnl
80r---------------~--------------------~---f r a
--c 60 f, t u r e 40 ~ a
m O~--~----~I----~----~--~~--~----~--~ o 0.2 0,4 0.6 0,8 , 1.2 1.4 1.6
distance betw6en blast holes m
Figure 8-12: Modulu.' verlu. 81 •• t Hole Spa~lng
100
m 801-o - ---d ~
u 1 60 ~ u . ,.. 8 • 40
G p 20 a
0 0 0.2 0.4 0.6 o.e , 1.2 ',4 1.6
distance between blast holes m
) 160
)
..•
Figur. '-13: Modulu. Y.rtu. Fr.ctur • ....,.n.lty
( ----------~-
190
m 80 0 d
~ .u , 1 60 ---u 8 -~. ....
40
G p 20 ~ 8 ..
0 \
0 20 40 60 ao fracture denslty per m
k (
c 161
o
(
f fracturing would be expected to occur midway between the blast
holes, however, a slight reduction in modulus may occur. The
trends of the relationships are obvious, however, fitting a curve
to the data is not easy due the hig~scatter and relatively small
data base. The information does provide a starting point for practical destress blast design.
8.3 BLAST VIBRATION STODIES
A nurnber of test blasts were conducted throughout the mine in an
attempt to establish attenuation characteristics of the rock
masse A blast vibration monitoring instrument and the mine's
microseismic monitoring equipment were used in the study.
Neither of these instruments was capable of providing information
on vibration within the fractured zone about the blast holes.
Peak particle velocity attenuatlon curves were developed for the
elastic response region beyond the area of fracturing. The
attenuation characteristics of particle velocity change
significantly once the stress wave is below the amplitude which
can fracture rock. In order to determine attenuation and
establish peak particle velocity as a predictive tool for rock
fracture, measurements must be taken in the region were
fracturation and other nonelastic processes are occurring.
Test blasts were conducted by drilling holes 0.8 m into the drift
wall. Holes were spaced from 3 to 28 m away from a cement pad
monitoring station were a three axis geophone was mounted. The
geophone was linked ta an Instantel Limited DS377 blast
vibration monitor. Limitations associated with the rsponse
frequency and sampling rate of the OS3?7 may have adve~sely
effected the results of the tests.
The 05377 has a maximum frequency response of 250 Hz, blasting
produces frequencies greater than this. These frequencies are
filtered out as the wave travels through the rock. Using the
DS377 close to a blast, where high frequencies still exist, will
162
(
... lead to erroneoUB re'sults aB the entïre wave form will not be
captured and analyzed. Data sampling rate is inversely
proportioanl to sampling time. A long monitoring period results
in a low samp~ng r~; this may introduce error. Typical output
from the ins~rJmen~s shown in Appendix 3. \ ''--
The OS377 proved to be very unreliable. In sorne instances the
tri~ger mechanism would fail, other times keyboard operation
would lock for no apparent reason. It is concluded that the
instrument is not suitable for near field monitoring and is not
rugged enough for underground operation.
Attempts were made to use the mine's microseism!c system linked
to a wave form analyzer to monitor blasting. Details of the
microseismic monitoring system are given in Chapter 6. Knowing
the response characteristics of the transducers and any signal
amplification, it is possible to determine a peak particle
velocity or acceleration (depending on type of transducer; ~
geophone or accelerometer). The study used aIl geophones with a
peak resRonse of 6.8 volts peak to zero; transducers within 100 m
of a blast were over-driven. Results from far field blasting
were highly variable indicating the non homogeneity of the mine
with respect to sei smic wave transmission.
Energy values generated by the energy integrator on the
microseismic system were investigated for possible attenuation
characterization. Results of both the Instantel and MP250
monitoring are presented in Table 8-3. Regression analysis on
the 05377 data resulted in a best fit curve of the form;
1
/ )'
/ . where R = radial distance from blast
w = weight of explosive ,1
1 i
The limitations of the data collection must be considered wh~n using this equation for blast design purposes.
163
/ l
)
l
3
\ "
0
~ Il ,"
Tabl. 8-3: R •• ultl of Blaltln. ~.It.
i
RECORDINGS FROM ELECTROLAB MP:50 AND INSTANTEl 05370
LOCATION EX PL WEIGHT ("g)
INSTANTEL DIST PPV
("''''/5 )
MP250 COMMENTS OIST ENERGY
( PI )
----------------------------------------------------------------------------
-- 1470 ANFO 2.35 12.00 7948 OESTRESS 8LA'il 1470 ANFb 2.35 13.00 5954 SUSPECT DATA 1470 ANFO 2.35 15.00 -. 9165 POS5IBLY TWO
1470 PMEX 0.30 11.00 6865 HOLE 15 TES18 1470 ANFÔ 1.88 -12.50 9020 HOLE #\ TEST9 1470 PMEX 1.50 12.40 8948 HOlE #2 TEST'" 1470 ANFO 1. 88 \1.90 9198 HOLE #3 TEST' , 1470 PME'X 0.30 11.90 8332 HOLE 13 TEST? 1470 PMEX 0.30 10.40 7292 HOlE #5 TES1:' \470 PMEX 0.30 10.40 8011 HOLE #5 TES T 5
1470 PMEX 0.30 44.20 5.40 14.40 9073 HOLE 84 rES TI
1470 PMEX 0.30 48.80 5.20 9.90 4706 HOlE #7 TEST4
1470 PME X 0.40 45.80 2.22 1470 PMEX 0.30 44.80 20.70
1604- ANFO 1. 92 61.00 5.41 1604 ANFO 1. 92 61.00 4.84 1604 ANFO \.92 9.20 35.58 1604 ANFO 1. 92 70.20 3.44 1604 ANFO 1. 92 6\ .00 4.00
2700 PMEX 0.40 4.60 21.24 2700 PMEX 0.20 3.10 14.56 2700 PME X 0.20 9.80 6.00 2700 PMEX 0.20 7.60 10.-42 2700 PMEX 0.20 7.60 11.00 2700 PMEX 0.20 9.80 8.08 2700 PME X 0.40 15.30 4.66 597.00 21 -
2700 PMEX 0.40 18.20 Il.22 597.40 59 2700 PMEX 0.40 21.40 6.S6 597.40 53 2700 PMEX 0.40 21.40 Il • J0 2700 PME X 0.40 3.10 29.58 596.60 99 2700 PMEX" 0.40 4.60 Il.38 . 596.60 91 2700 PMEX 0.40 6.10 Il.02 596.60 90 2700 PME X 0.40 9.20 4.12 597.00 89 2700 PMEX 0.40 12.20 7.46 597.40 113 2700 Pf1EX 0.40 24.40 2.80 597.40 69 27.00 PMEX O.40 27.50 3.84 597.0O 108
'2700 Pf1EX 0.40 27.50 3.64 597.O0 107
164
•
•
c.
8.4 SUMMÂRY AND CONCLUSIONS OP BLAST PRACTURING STUDIES
~ The studies, reported on in this chapter, allow a number of \ conclusions to be made as to the extent of fractures existing
around a stope, and the extent of fractures induced by confined
blasts. Modification of rock mass properties by blast fracturing
is also quantified.
stope pillars generally have an intact core, while, a fractured
zone borders any excavation openings. The fracture zone in the , . l670'W A pillar waS measured to extend a distance equal to 18
-percent of the width of the excavation opening. Fàctors such as
stress level, excavation shape and pre-existing discontinuities
will also affect the depth of the fractured zone. Fractures
caused by excavation damage or destress blasting forro
predominately alqng pre-existing discontinuities. Destress
blasting should not be conducted in the fractured zone,
surrounding a pillar, as it may result in cratering, as has been
experienced at other mines, Chapter 4.
The CRLM andesite was observed to be highly variable in
mineralogical composition and fabric, however, its mechanical
properties were found to be relatively uniforme
Blasting AN/FO in confined 45 mm holes resulted in detectable
changes in the modulus of deformation, midway between blast holes
wh en blast holes were spa?ed up to 1.6 m apart. Modulus of
deformation was measured in situ with a Probex-1 bore hole
dilatometer. Fractures, detected by impression packing, were
observed when blast holes were spaced up to 1.4 m apart. The
trial blasts indicated that fracture frequency 'decreases with
increasing blast hole spacing, modulus of deformation decreases
with decreasing blast hole spacing, and modulus of deformation
decreases with increasing fracture frequency. The relationship
.between these parameters is approximately linear.
165
'.
o
Damage beyond a line drawn perpendicular to the end of the
explosive, charge is minimal in confined blasts. This ia
attributed to a reduction in stress supurposition beyond the end "'of the charged hole.
In order for destressing to be successful, fracturing must oceur
across the plane connecting adjacent blast holes. A wide blast
hole spacing may result in an unfractured zone between holes.
The unfractured zone wi~l attract stress and may result in a rock
burst at a latter time. In CRLM andesite the maximum hole
spacing should be 1.6 m for 45 mm blast holes using AN/FO.
Greater fracturing and rock mass modification would be achieved
with a tighter spacing.
Neither the microseismic system nor an Instantel-DS377 blast
vibration monitor were capable of determiqing peak particle
velocity characteristics of the blast wave for analysis of
fracturation. Analysis of fracturation requires knowledge of r
blast wave behavior in the non elastic response zone surrounding
the blast hol:" A peak\particle velocity attenuation curve for
sefsmic wave~ traveling beyond the blast fractured zone was
determined, however the supporting data is hi~hly scattered.
166 1
•
•
•
, DIBTRIBS BLAST DISIGN
Destress blasting has been shown to be an effective tool for the
control of rockbursts. However, without proper design, a
destress blast may be ineffective or result in other~ (
potentially more severe, problems. This chapter presents
guidelines for de stress blast design based upon the research and
literature review conducted for this project. The case being considered is de stress blasting .in andesitic rock at Campbell Red . '
~ake Mine (CRLM). with appropriate modification the design
criteria are applicable for confined blast fracturi~ in any
situation. o ,
The design of blast patterns for destressing has generally
proceeded in a haphazard fashion with very little scientific
investigation. Further more, this author has observed that many
modifications are made to existing destress blast design for no
apparent technical reason. Blasting patterns ~eveloped at other mines, are often used as a starting design. A trial and error
approach, as is often used for optimizing production blasts
designs, is not suitable because of the small number of destress
blasts norrnally conducted. The large nurnber of influential
variab~es, and the difficulty of assessing thern adds to the
problem of destress blast design optimization.
Destress blast optimization i5 considered achieved when the
modulus of deforrnation is redu'ced to the desired level by
interconnecting fractures between blast holes. . Fradturing
should not be so extensive as to impede subsequept mining,
although this criteria may not be possible if the desired modulus or strength reduction is great .
167
o
o
9.1 BLAS~ BOLE SPACINr AND GBOMBTRY ./
Blast hole spacing, and to a certain extent, hole diameter are
the most important variables in d~stress blast design. Hole spacing can be adjusted to achieve the desired degree of fracturing regardless of other parameters and in situ conditions, The appropriate blast hole spacing can be determined by powder factor calculation, peak particle velocity analysis or experimental investigation. AlI these methods are derived from
observations of fracturation under a given set of conditions:
Should the in situ environment change, the spacing selection may
no longer be valide
Destress blasting of stope pillars at CRLM has used powder factors between 0.2 and 0.7 kg/m3 . A review of these blasts
reveals that there were no significant differences in damage
caused by using different powder factors. The amount of energy
imparted to the rock is dependent upon \the chemical energy of the
explosive, the hole diameter and the rock - explosive
interaction. Powder factor considers the energy imparted to the rock as being dependent upon the weight of explosive only.
Destressing usually involves blasting only a single row of holes along the length 9f a pillar. The width of the pl1lar, which la
used in the powder factor calculation, la of no relevance.
powder factor ia a simplistic blast descriptor and is not
suitable for destress blast design.
Peak particle velocity (PPV) is related to strain which in turn
ls" related to fracture extension. The PPV requited to lnduce
fractures can be calculated theoreti~ally after determining the
strain at- whic~ the rock fractures. A PPV of 1000 mm/a ia <
generally accepted as sufficient to create fr~cturea in competent
rock. An estimate of PPV is given by the equation
;168
' , ,
()
.,
(
, '
.'
.... ' ,
" wO'~ ) -n PPV = k R / ,.
where k = <J:round transmission constant n = empirical constant based on geology and explosive .. R = radial distance w = weight of explosive detonated
\t 'Knowing values for the k and n, the above equation can be rewor~ed to determine the desired weight of explosive to achieve the peak particle'velocity requlred for fracture extension to a -
specified distance. Values of k and n determined béyond the imediate vicinity of the blast hole cannot be used as they do not ~eflect the high stress conditions under which fracturing 'occurs.
~ , Determination of fracturing by peak particle velocity is not recommended beca~se of the inaccurate explosive energy considerations (based on weight of explosiv~J and the difficulty of.determining the transmission constant in the vicinity of the blast hole where fracturing is taking place.
),
In situhmeqsurement of fractures and the corresponding reduction in modulus between blast holes provides the most accurate inforrnat~on,for destress blast design. Figures 9-1 and 9-2 present results from in situ tests conducted i~ CRLM andesite.
·'Best fit curve~ were determined to be linear, these are also plotted on the' Figures. 'The equation Of the curve relating
" distance between -blast holes to fracture intensity at the mid point is:
frqcture density = 117 - 83.3'* (distance betweep blast holes)
The equation rel~ing dilatometer modulus of deforrnation to fracturation is:
modulus of defo.rrnation = 64.2 - 0.5 \ (fracture density~
~he corresponding corre~ation coefficients are 0.62 and 0.59.' Although low/~hey are not unreasonable when dealing with
169
•
0 l
,--
0
" .
,
Figure 9-1: Relatlon.hlp aetween Fracture Den.lt, and ala.t Ho" Spa01n1
f r El 80 c t u r
60. e
d e -n 8
40 1 t Y
20
" p e 0 r
0 0.2 ; 0.4 0.6 0.8 1 1.2 1.4 '.6 m distance between blast holes m
Figure 8-2: Relatlon.hlp aetw.en Deformation ~odulu. and Fracture Oenelty
• 1'00
m 80 0 d u 1 _80 u e
40
-G p
20 a
0 0 20 40 60 80
, fracture denslty "par m
,1' i
170
•
(
c
/
/
blasting studies where scatter is typically very high •
These studies were conducted in the unfavorable situation
requiring fracture extension perpendicular to the major ',-
discontinuity direction, as su ch they are considered
conservative. The reduction in modulus of deformation and reduction in~strength required for a successful destress blast is
beèt determined by numerical modeling. •
It may ai first seem logic~l to use a small number of large
diameter holes to achieve the desired fr-acturation. Larger holes
result in quicker detonation velocities, and for the most part, \
greater detonation and 'explosive pressures. Although not we,l,!
defined, there is a limit beyond which increasing the detonation
energy released from a single hole results in no increase in the
the extension ef fractures; the only result is an in crea se in
the crushed zone about the blast hole periphery. Further work is
required to determine the optimum hale diameter. The area about
thedblast hale has the greatest concentration of fractures 1
running in aIl directions. This zone causes the greatest
alteration ta rock mass properties. The use of small holes is
desirable in t:hat they result in a greater number of these areas
of high fracture concentration.
Provided that a sufficient number of fractures are created, then
the width of the fracture zone is not important, however, the
orientàtion of the fractures is consequential. Fractures created
perpendicular to the principal stress direction do little ta
reduce the modulus or strength. Fractures created at angles
other than 0 or 90 degrees to the principal stress direction have
much greater effect. Madel studies have indicated that fractures
at angles of 60 or 30 ~egrees ta the direction of principal
stress cause the greatest reduction in modulus of deformation and
uniaxial strength. An arrangement of two rows of holes offset 50
that connecting lines form a 60 degree angle with the direction
of principal stress should result in the gre~test decrease in '\
modulus of deformation.
('
171
•
•
9.2 EXPLOSIVE SELECTION
The effectiveness of any explosi~ depends on both the explosive
properties and properties of the ~aterial belng blasted. Ground
response to various explosives can be estimated by numerical
means or investigated by in situ studies. In situ studies
provide the rnost reliable assessment of which explosives are
suitable for destress blasting.
w The total energy content of ~n explosive is readily calculated by .
thermodynarnic analysis, however, these calculations assume the
explosive to'behave in an ideal manor, and do not consider th~
influence of confinement and hole diameter. More important is
the rate of energy release and the energy division between.
detonation pressure, explosive pressure and wasted energy. It
should be noted that the energy division i5 a function of many
parameters including density and hole diarneter. The rate of
energy release can be significantly changed by small changes in
borehole diameter and density, although the degree of change ls
different for different explosives.
It is desired that the explosive induce a minimum number of
fractures extending at least to the midpoint between blast holes.
Fracturation must be extensive enough to reduce the modulus of
deformation, and rock strength, to the desired values. Sorne
researchers, e.g. Favreaù (1986), consider the detonation
pressure to be insignificant in creating fractures in a confined
environment with no free face. This means the explosive gas
pressure must exceed the rock or discontinuity strength as weIl
as the in situ stress across the desired fracture plane. This
author believes the explosive gas pressure is responsible for
sorne fracture creation through wedge action, however, with in
situ .stresses exceeding 50 MPa it seems unlikely that it is the
main mechanism .
The detonation pressure is of a much greater intensity, although
of shorter duration, than the explosive gas pressure. The
172
•
~'
~
c
«
J
explosive pressure generates radial and tanqential seismic waves
,with compressive and tensile components. In a high in situ stress regime, where significant stress superposition may
occur,blasting may lead to compressive failure out to a
considerable distance from the-blast hole. Tensile failure is
unlikely to occur except in the immediate vicinity of the
borehole and in the direction perpendicular to the desired'
fracture direction. This discussion has assumed the area to be
fractured is already highly stressed. In preconditioning a high
stress environment may not exist; in this case the tensile
component and action of the gas pressure would be much more
significant.
The greatest detonation pressure occurs with the most violent
detonation reaction. A measure of detonation reaction is the
velocity of detonation. High velocity explosives are therefore
preferred over lower velocity types. There may be a limit beyJnd
which increasing the detonation pressure results in an increased
cru~Q~d-4Qne, but no increase in fracture extension at greater
distances. "~ore work is required to investigate the shock wave
phenomena in rock, and this particular problem.
Factors that must also be considered in explosive selection are
availability, cost and safety. It may be cheaper and more
, practical to use a regular production explosive than to use a
specialty product. Additional drilling costs required, in order
~o achieve the desired fracturing with production explosives, may
be less than the cost of buying and stocking specialty products.
It mu}t be considered that no savings will be realized if the
blast is not successful because of poor explosive selection, or ~
~ for any other reason.
9.3 OTHER BLAST DESIGN CONSIDERATIONS
There are a number of additional factors influencing blast
results. ~hes~ include the location of the holes in relation
173
l
•
to the ore zone, the detonatibn sequence, mode of detonation, and
the use of stemming. r
1 Little information is available on the effect of destrè~
blasting on and off vein. Karwoski et al. (1979) conducted pre
conditioning blasts on and off vein, they concluded that there
was no difference in terms of subsequent ground control problems.
Their computer modeling suggested that on-vein blasting is
preferential. The advantage of off-vein blasting is that
subsequent mining need not be conducted in fractured ground, nor
is there a risk of mining into residual explosives. Further
numerical and in situ investigations are required before any
conclusions can be drawn.
The greatest benefits of blast hale interaction are achieved when
\ detonation Gccurs simultaneously with the passage of the
C) detonation pressure wave from the adjacent hole. This can be
achieved by delay timing or using pressure sensitive caps. The
safety hazard of pressure sensltive caps makes them undesirable.
High quality mili-second delay caps must be used as conventional ~
caps have a high time scatter. Blasting simultaneously, using \
conventional cap, with equal delay may provide good results, the
delay scatter will result in some constructive stress interaction.
When blasting on-vein it is important to ensure complete
detonation to eliminate the possibility of mining into residual . explosives. Multiple initiation points are recommended. Holes
should not be traced with detonating cord as this may initiate
undesired deflagration. Adequate priming should be used to
ensure the explosive reaches the maximum detonation velocity as
early as possible.
The contribution of the explosive gas pressure ta destress
blasting i5 not weIl defined. Any contribution should be
enhanced by maintaining the gas pressure for as long a time
period as possible. stemming should be used for this purpose.
The length of stemming should be the entire uncharged collar.
174
•
•
•
The length of hole not charged depends upon the condition oÏ the
rock. A highly fractured zone surrounding the collar will lead
to large craters if an insufficient uncharged l~ngth is not l~ft.
9.4 NUMERICAL MODELING FOR DESTRESS BLAST DESIGN
~
Nurnerical rnodeling, run on comp~ters, are use fui for analysis of
a number of aspects related ta destress blast design. The
accuracy of the results of numerical simulations must always be
consideredi this is dependent upon the sophistication of the cgde
and the accuracy of the input parameters. Any model must be
calibrated by back analysis, and shouid be verified against in
situ conditions whcnever po',~Jble. Nuwericai modellng for
fracture prediction is not yet ta astate where it can reliably
model the effects of in situ stress and pre-existing
discontinuities. It is hoped that more sophisticated models will
become availâble in the near future .
Numerical modeling of stresses is a weIl accepted and proven
technique. The method is extremely usefai for determining what
rnodulus and strength reductions are required for a particular
destress blast ta be successful. The models can also be used to
10cate areas requiring destressing, what aerial extent should be
fractured, and the potential for creating new areas of
instability as a result of stress transfer .
175
J
10 CONCLUSIONS
The technique of destress blasting for rockburst control has been utilized with apparent success at Campbell Red Lake Mine.
However, the mine recognizes that the need exists to imp~ove the understanding of rock mass response to blasting~ upon which
dastress blast design and optimization are based. Without proper consideration and design, a destress blast may be ineffective, or
result in other, potentially more severe problems. In situ
studies and a literature review, aimed to provide a better
understanding of rock fracturing and develop destress blast
design criteria were carried out during the course of this
project. Conclusions of the study are presented in this chapter.
'J . The nature of the fracturing ptoduced by explosives i8 dependent
upon bath explosive and ~ock properties. The literature review
and experimental studies suggest that the most influential
parameters are discontinuity strength, discontinuity orientation,
existing in si tu stress, and degree of confinement. "
" Given a blast hole size and explosive type, the blast hole
spacing can be adjusted to achieve any degree of fracturing , between blast holes. The extent of fractures will be
gradational; the most extensive fracturing being nearest to the
blast holes. The extent of fracturing ~t the midpoint between
blast hales was found ta be linearly, inverse-proportional ta the
blast hale spacing.
Fracturing alters rock mass properties. For the purpose of
destressing it is des~abie to reduce po~h the rock mass modulus of deformation and st'rengthi this is achieved by introducing
fractures in to the rock. The modulus of deformation was found
to be linearly inverse-proportional to fracture frequency. It
was observed that peak, unconfined compressive strength decreased
with increasing fracture frequèncy, however, insufficient data
was collected to allow for any statistical relations~ip to be
determined. .. \ l
176 (
(~
(~
(
Blast fracturing studies were conducted in Campbell Red Lake Mine andesite, using ammonium nitrate and fuel oil explosive detanated
in 33 and 45 mm diameter holes. Rock mass canditioning,
resulting from fracturing, was measured at the midpaint between
adjacent 45 mm holes. Using an impression packer, fractures were
abserved where blast hole spacing was a maximum of 1.4 m. Using
a borehole dilatometer, a reduction in modulus of deformatian was
abserved where blast hale spacing was a maximum af 1.6 ID.
Blast fractur.1pg, araund single 33 mm diameter blast hales, was • ..1
obverved to extend upwards of 0.5 m along pre-existing
discontinuities. A radial, heavily fractured zone extended
between 10 and 100 mm about the blast holes. A crushed zoner
which was independent of pre-existing discontinuities extended
between 2 and 30 mm about the blast hales. AlI fracturing, with
the exception of the immediate blast hole periphery, occurred
alang pre-existinq discontinuities.
Blast in~uced fractures were observed to not extend significantly
past a line drawn perpendicular to the end of the explosive
charge. This ls a result af a decrease in constructive stress
interaction. In an area close ta the cent pr of an explasi ve
charge, the cumulative effects of energy being released along the
entire length af the charge must be cansidered.
Studies in narrow veln stopes at CRLM provided no evidence that
single hale destressing is an effective means af rockburst
control, however, it is possible that benificial changes, beyond
the detection af the monitoring equipment, occur. Destressing
was conducted using a single, 4.5 m long, destress hole in a 3 m
breast round. Stope ,monitoring was conducted with a microseismic
system, e~tensometer, vibrating wire stress gauges, visual
inspections and fracture mapping. The study revealed no
significant differences, relating to a reduction in rockburst
potential, between breast rounds with a destress hole and those
without. It should be noted that, as a safety precaution, CRLM
continues te use this destressing methed.
177
.. <,
J
Destress blast design requires the knowledge of what level of
modulus and strength reduction is required, what amount of
fracturing will cause the desired alteration to the rock mass
properti~s, and what blast design will achieve the required
fracturing. The level of modulus and strength reduction,
required for the destress blast to be successful is best
deterrnined by numerical modeling. Nurnerical rnodeling ls also
useful for determining what area of ground requires destressing,
and if the destressing is likely to cause unstable conditions in
adjacent areas. The relationship between modulus of deformation
and fracturing is dependent upon rock type, in situ stress and
fracture orientation. The rnost accurate method of determining
the forrn of this relationship is by in situ studies. The blast
design required to aehieve a specifie amount of fraeturing can be
deterrnined by predictive methods or in situ studies. Predictive
methods, either numerieal or ernpirical, are at present not v~ry
accurate for frac tu ring caused by destress blast~ng; this is a
result of insufficient consideration of in situ stress, pre
existing discontinuities and blast confinement. The relationship
between blast design and fracturing is best determined by in situ
studies.
A number of additional factors pertaining to destress blast
design were reviewed, including; explosive selection, blast hale
geometry and initiation sequence. An explosive with high
velocity of de~onation is most desirable.
the greatest detonation and gas pressures.
These explosives have
There is a limit
beyond Vhich increasing the explosive energy imparted to the rock
is no longer beneficial. Beyond this limit the energy serves
only to increase the crushed zone. The most desirable blast hole
configuration, for reducing modulus and strength, is two
staggered lines where a line drawn between holes makes a 60
degree angle with the normal of the hale lines. Explosive
initiation sequence should be designed te enhance stress, wave ~
interaction. This carrbe achieved using pressure sensitive caps
or with low delay scatter caps.
178
~
, t ; ~" , _
APPENDIX 1 Thin Section Descriptions
f /' , /
c , .
•
"
,
179 :.-,
)
o
This appendix contains selected thin section descriptions for both. the ~ 6~4 'EW stope crown pillar and 1470 drift. The sampl,es are ~dent~f~ed by a sample number followed by the location. Samples from the 1604 EW stope are listed first, they have a sample number suffixed with either a 400 or 401. Samples from the 1470 drift are prefixed with a 4. Thin section photographs
-are located at the end of the appendix.
- -- 1
2-400: located 4.1 m from hole collar, 4.1 m from 1604 destress holes. ~he core was logged as an&esite, blue-medium
Mode: biotite 1-2 % calcite 15-20 % chlorite 30-35 % fine matrix 30 % muscovite opaques 8-10 % quartz 5 % plagioclase
The absence of calcite is expected since the sample is not in an area of hydrothermal solution pa$sage. One small quartz-calcite vein does cut the section. Coarse, anhedral q~artz crystals are present, these may be relie or possibly vein rnaterial. The absence of biotite and high percentage of chlorite indicate extensive low grade metamorphism of the ferro-magnesium mineraIs. No fracturing or other evidence of stress is observed in this sample.
3-400: located 8.2 rn from collar, between 1604 destress holes. The core was logged as andesit~, blue-medium.
Mode: biotite 5-6 % Calcite )25-30 % chlorite 25-30 % fine matrix 40-45 % muscovite 8-9 % opaques 5-6 %
l quartz plagioclase
A number of small veins eut this sample: the veins contain quartz, calcite, mica and plagioclase. The presence of biotite -indicates less low-grade metamorphism than sample 2-400. The high calcite percent age indicates close proxirnity to hydrothermal conduits. No fracturing or other evidence of stress is observed in this sample.
1 4-400: located 12.2 m from collar, 4 rn from 1604 destress noles. Sample i5 above pillar separating 1604 A and B veins. The core was, logged as andesite, green-medium, with tourmaline faults.
180
'C
(
Theo sample is divided into three distinct zones; a zone of coarse grained qua~tz, calcite and plagioclase, an zone of fine matrix,
. and a zone of coarse quartz with associated unknown porphyroblasts. The unknown porphyroblasts arè prismatic with
· polysynthetic twining, extinction is parallel; 1-2 degrees.
5-401: located 12.5 m from collar, between 1604 destress holes. The core was logged as blue-medium andesite with smooth fIat joints and parallel foliation 75 degrees to core axis.
Mode: biotite 10-15 % ~ calcite 30-35 %
chlorite 10-12 % fine matrix 35-40 % muscovite 5-6 % opaques 5-6 % quartz plagioclase q!V
A number of quartz-calcite-plagioclase veins cross eut the sample. The chlorite was often found as anhedral blades enclosed in calcite. Calcite and plagioclase trystals contained intracrystal microfractures. The fractures were anastimosing and in sorne instances the fractures fol'owed a curved path. The fractures were limited to isolated areas of the sample. Quartz associated with fractured crystals showed no fracturing, but did exhibit undulose extinction. Undulose extinction indicates high stress although it is usually thought to be produced by the in situ stress field, not blasting. The sample contained_ deformed mica cry~t~s and cUFved twin lameli in the calcite, both result from high in situ stresses.
4-30: located 0.3 m from collar, midway between 3 m spaced destress holes. The core was logged as andesitè with quartzcalcite infilled joints parallel to cleavage, 6 cm spacing. The core was fractured along infilled joints. The core is quite competent, RQD is 95 percent and core recovery is 1~ percent.
Mode: biotite 10-15 % ( calcite 30-35 % chlorite 25-30 % fine matrix 30-35 % muscovite 1-2 %
181
opaques 4-5 % quartz plagioclase
, The sample contains many calcite flasers and augens in a chlorite matrix,. as weIl as sorne chlorite augens in a calcite' matrix. Fo+iation is strongly developed, and ls post contact metamorphism (hydrothermal intrusion), as indicated by the flasers and augens. ' See Figure A)-l.
4-120: located 1.2 m from collar, between 1.3 ~'spaced destress holes. The core was logged as andesite, blue-medium, with quartz-calcite filled joints, mostly paraI leI to cleavage, 4 cm spacing. Fracturing was along cleavage in rnost cases,-a few new fractures ran 015/85. RQD of core was a, core recovery was 90 percent.
Mode: biotite 6-7 % calcite 35-40 % chlorite 20-25 % fine matrix 20-25 % muscovite 1-2 % opaques 3-4 % quartz plagioclase
D
Good foliation development, weak augen development.< .
4-160: located 1:6 m from collar, between ).3 rn spaced destress holes. Logged as andesite, blue-medium. The sectiun bf core contained 0~005-0.1~m dises, indicating high stresses. Core RQD was a, core recover~ 94 percent.
Mode: biotite 6-7 % calcite 35-40 %
'chlorite 20-25 % fine matrix 20-25 % muscovite 1-2 % opaques 3-4 % quartz
1 plagioclase
i
The sample was very fine grained comparéd to others,. If stress induced features were created it is un1ikely,that they would be visible. The foliation is strongly developed and üefined b~ chlorite ribbons. WeIl developed flasers and augens ~re a1so present. Biotite also i5 in ribbon forme See Figure Al-2.'
4-190: located 1.9 m from co1lar, between 1.3 rn spaced destress hales. Core logged as andesite, blue-medium, the core wàs" broken up, maximum size 0.05 m. Sorne infilled joints were evident . Core ROD was 0,' percent recovery was not determ,ined.
, ,
182 \
'" ...
j
~ . )
o li '
Mode: biotite 10-12 % calcite 40-45 %
.' chlorite 20-25 % fine matrix 20-25 % muscovite 2-3 % opaques 3-4 % quartz pla~oclase
Weakly developed foliation.
~~210: located 2.1 m from collar, between 1.3 m spaced destress holes. Logged as andesite, core was fractured along clea~age,
':"
strike 130, dip 80 west. The fractures were not infilled. \ Quartz calcite infilled.jo~nts parallel cleavage, 3 cm spacing, \ minor infilled joints 100/80, 5 cm spacing. Core ROD was 0, recovery was 100 percent.
Mode: biotite 10-12 % calcite 40-45 % . ~ chlorite 15-20 % fine matrix 20-25 %
'f
muscovite 2-3 % "- opaques 3-4 %
quartz plagioclase
Strongly developed foliation, chlorite ribbons. Very fine grained, similar to 4-160. Sorne coarse grained calcite aggregates, weakly developed auge~ and flasers.
~ , r-4-284: located 2.84 m from collar, between 1.3 m spaced dlstress holes. Core logged as andesite, dark blue. Quartz calcite infilled joints parallel. cleavage, 8 cm spacing, mi~or infilled joints at 040/30, lb cm spacing. The core fracturèd perpendicular to the core axis. The 'core also was subject to high drill wear indicating a softer rock. ROD and CQre recovery were both 100 percent.
Mode:
.
biot~e 12-15 % calcite 40-45 % chlorite 16-20 % fine matrix 20-25 % muscovite 1-2 % opaques 4-5- % quartz plagioclase
The sample contained cross c~tting_calcite veins. Broken veins form augens. Sorne chlorite ribbons are present, foliation not strong. Fine grained, similar to 4-160 and 4-210. The hiqh biotite content accounts for the darker color and possibly tor the observed drill wear. ~
183
(
,J
(
" •
1 r ,. ,
4-325: locat'ed 3.25 m from collar, between 1.3 m spaced <1estress holes. Con! logqed as àndesite, blue-medium. Vis.ible cleavage strike 130, dip 80 west: No fractures. Quartz calcite filled joints paraI leI cleavage, 6 cm spa~ing. The sample·is from- a location remote from theodestress blasts. Core RQD and ~ecovery were both 100 percent.
Mode: biotite 5-6 % . calcite 40-45 % • chlorite 29-25 % fine matrix 20-25 % muscovite 4-5 % opaques 3-4 %
~ quartz , plagioclase ----....:.
One large quartz-calcite .vein cuts the sample. Muscovite i5 developed as porphroblasts crystaîs within the fine matrix. Good foliation. The muscovite cross cuts the foliation, indicating a second'metamorphic'event. See Figure Al-3. ~
\
..
1 184
'. 1
, .
•
o·
'r
Figure-Al-l: Sample 4-30, Chloritic Augen in Calcite Matrix.
1 .-
• Figure Al-2: Sample 4-160, Strong FOliation, Chlorite Ribbons,
Well Developed Augens
•
185
••
c
(
)
Figure Al-3: Sample 4~5, Muscovite Porphryoblast \
186
. \ • Î
0-
APPENDIX 2 Bor. Hoi. Impr ••• lon Mapa
. ,0
.. , -
"
p "'l;..,.'
o t
187
ho •• 1
c colla" , \ --1-
f -\
) r ..,.,-
'-../
'-
/'-..... - ) /
10 cm ) '---'
"
, ,
- ! , !
1
/ t 1 ,
.- 1 /,
/ ,
1
~
t 1 ,
J c i ,. to.
"1
1
hol.2
• coll~u -'
• t •
6 1~~" ~'
"
S:·
L '\
l,
l,
'1. l"'.
, - .>.
10 cm 1..-.-1
"
..
l
o to.
189
.J.J,
\' ,.., '- J'
" " ~ hol. 3 .
"'. I~
• L ~
. "
. .... .
,
"
\ •
\
-. Aï ----:------::------:---~-----:--~----_._._-_ .. __ .... -'f _,'
\ ' "
, ~ .
,/
\ ~ , /
1 (
1
\-
o
: __ --+-__ ........ t~.
[ .
o ,(
191
,\
hol ...
coUa'
c --
-,
<
'/
,
10 cm -
( \
L -'\
'-". 1
• r ~.
.....J-...
/ -- ~
'-
--... ~
- -
1
,~
" \ ...
-D
\.., \-~
-Ar '\ r
J ....
0 -
t
, .
to
. 193
- . .... ..-J
..
c' •
APPENDIX 3 INSTANTEL DS3700 Output
...
194
•
•
D
\ '
o
_:. .... ~1_,.~._ .;, .... ':sMI-
~.; j: :=. cCI ........ dL,:"g SC"RCf
- ~ 4'",:8 l FUEl i=:';;'"(1 T lnf
-1--
Î
INSTANTEl OS-3n
.l)':;)
ll'll" Cl.!Ll-EN "CGIU llNluE-RSITT l,pa DR. 1 • n/5
Ge ophOllQ
13 e7 • n/, J Socond.'
:zs ,"'v,
m .. .. .. Xl .. n ., ...
.\----~------~--------~~--~~~ . ,
•
115
..
........... " ...... _ .... '.' .~.,
• ........ _ ....... ~ ....... .
l
'1
.--l-.~~ _~ __ , •• l .. _._ ~
CHECK TR~NSVERS~L TE$' R:~_ ;_1 F T.~C or -20. Fu·~a OV'2, ;", '
Ir ,g9.,.d go
euk Se!
LOn9. Ch. "'
9)' 1 ~ ~8
Fob :6, ~~
13.1007 Se:
9."1301' PI ' 42.15 P.-/_--
GIPRbQ~o , • U!II!S /nc'" • .:;. 4, 7j] "J~/ ..
112 08 8 P •• • 0'1110 p,
etH Ueleeit ••• ;
Tf .n ..... ' ••
U.rlIC.'
\.on, 1 t "d.
.'98t> 11~90 .11152 4.211 .411'4 HI. 49
.118 28.'4
."!h •• , ..
~.~.i. c:
: .. :. .... , ,.. 'i.e:.
.l''!''.',,,, ., .. '" i.:
}r:",,""'", , • .l'.c
..
\ ..
J
(.
(
(
UJ'ZUNCZS
Amadei, B., 1985. "The Influen.ce of Rock Mass Fractuf.ing on the Measurement of D~ormability by Borehole Expansion Tests", Proc. 26th U.S. Symp. on Rock Mech., Rapid City. pp. ~59-867.
Arjang B., 1986. "Field and Stress Determination at Campbell Red Lake Mine, Ontario", CANMET M&ET/MRL 86-57 TR.
Atchison T.C., Duval W.L, Pug'liese J.M., 1963. "Effect of Decoupling on Explosive Generated Strain Pulses in Rock Il •
USBM RI 6333 Blake W., 1984. "Rock precondition±n2L as a Seismic Control
Measure". Pr9c. 1st Int. Congo 9n ~ckbursts and Seismicity in Mines, SAIMM, Johannesburg. pp. 225-230.
Blake W., 1982. "Microseismic Applications for Mining: A Practical Guide". USBM Contract Report. No. J02J.5002, 1982.
Blake W., 1982b. "Destressing To Control Rockbursts Il. undergroupd~ Mining Methods Handbook, chap. 7, W.A. Hustrulid (ed), Society of Mining Engineers AIMM, New York, pp. 1535-1539.
Blake W., 1981. "Preconditioning an Entire Stope Block For Rock Burst Controi ll
• USBM OFR-51-81. f Blake W., 1972 IIRockbur~t Mechanics", Quarterly, Colorado School
of Mines. Vol. 67, No. 1. Blake W., 1972b. '''Oestressing Test at the Galena 1:1.iz:1e, Wallace,
Idaho". Trans. SME-AIME, Vol. 252. pp. 294-299~ Brady B., Brown E., 1985. "Rock Mechanics For Underground Mining"
George Allen and Unwin, London. ; Board M., Fairhurst d., 1983."Rockburst Control Through
Destressing: A Case Example". Proc. Rockburst Prediction And Control, IMM, London. pp. 91-102.
Butkovich T.,R., Hearst J.,R., 1976. "Prediction and Determination of Explosive Induced Fracture", Proc 2nd Conf. ~ on Expl. and Blast. Tech. pp 301-340
Church R., 1985. "Campbell Red Lake Mine Mining Brief: Geology". Report presented to Ground Control and Emergency Preparedness Committee.
Cook M.A., Cook V.O., Clay R.B., Keyes R.T., Udy L.L., 1963. "Behavior of Rock, During Blasting". Trans. 5MB. pp 383-392.
Cook J.F., Bruce O., 1983. "Rockburst at Macassa Mine and the Kirkland Lake Mining Area". Proc.' Rockburst Prediction and Control, IMM, London. pp. 81-90.
Crouch S., 1974. IIAnalysis of Rockbursts In Cut an~Fill Stopes". Trans. AIME, Vol. 256. pp. 298-303
Dickhout M.,1963. "Ground Control at the Creighton Mine of INCa Il , Proc. Cano Rock Mech. Symp., McGi11 Univ. pp. 121-139.
Du Pont, 1977 "Blasters' Handbook". E.I. du Pont de Nemours and Co., Wilmington, U.S.A.
Duva'l W.I., Stephenson D.E., 1965. "Seismic Energy Available From Rockbursts and Underground Explosions". Trans. SME. pp. 235-240.
Electrolab "250MP Microseismic System Specifications and Operation". Electro-Lab, Spokane, Washington.
Favreau R.F.", 1969. "Generation of Strain Waves in Rock by an ExplosioI1" in a Spherical Cavity". Jour. Geoph. Research. Vol. 74, No. 17. pp. 4267-4280.
196
•
o
\
( •
~ntzos O., 1980. "Destressing Practices in Ontario Mines". Memorandum, Ontario Ministry of Labour.
Hanson D. 1 Quesnel W. 1 ~ong R., 1981. "Destressing A Rockburst -Prone Crown Pillar - Macassa Mine" CANMET MRL 87- (TR) t
Harling J., 1965. Lakeshore Mine InternaI Report. <)
Hardy, R., 1982. "Applications of Aeoustic Emission Technique to Rock and Rock Structures: A State-of-the-Art Review". Acoustic gmissions in Geotechnical Engineering Practice. American
. Society for Testing and Materiala. pp. 4-92. Hedley D., 1987. "Rockburst Mechanies". CANMET MRL 87- (TR) Hedley D., Neumann M., Makueh A., Blake W., 1985. "Rockbursts at
Campbell Red Lake Mine"', CANMET, MRP/MRL 85 (OP). Hedley D., Bharti S., Blake W., 1984. "Fault-Slip Roekburst -at
the Falconbridge Mine". C~MET, MRP/MRL 85-115 (OP, J). Hill F., Plewman R.,1957. "Destressing: A Means of Ameliorating
Rockburst Conditions. Part II: Irnplementing Destressing with a Discussion on the Results so far Obtained". Jour. South African Inst. Min. Met. Vol. 56, 1957. pp. 270-289. ~
Hoek, E., Brown E.T., 1980, "Underground Excavations Jn Rock", IMM, London. .
Hustrulid A., Hustrulid W., 1972, "Development of a Borehole Deviee to determine the Modulu~ of Rigidity of Coal Measures Rocks", USBM Final Contract Report No. H010705.
Hustrulid W.A., Schrauf T.W., 1979, "Borehole Modulus Measurements with the CSM Celi in Swedish Granite", Proc. 20th U.S. Symp. on Rock Meeh., Austin. pp. 509-515.
ISRM, 1987 (International Society of Rock Mechanics) "Suggested Methods for Deformability Determination using a Flexible Dilatometer", Int. J. Rock Mech. Min. Sei. & Geomeeh Abst., Pergarnon Press. pp. 125-134.
Issaacson E., 1962. "Rock Pressure in Mines". Mining Publications Ltd., London.
Jaeger J.C., Cook N.G. W., 1976, "Fundamentals of Rock Mechanies", Chapman and Hall, London.
Jianyun M., Jiayou L., 1987. "The Phenomena, Prediction And Control Of Rockbursts In China". Proe. 6th Int. Cong. On Rock Mech., Montreal. AA Balkema, Boston. pp. 1135-1140.
~a shiro J.K., Harding R.C., Johannesson P., Korbin G.E.,
987, "Comparison of Modulus ValueS obtained frorn Dilatometer esting with Downhole Seisrnic Surveys and Unconfined
Compressive Tests at McKays Point Dam Site, California", Proc. 28th·U.S. Symp. on Rock Mech.,", Tuscon. AA Balkema, BostoI). pp.. 211-222.
Karwoski W., McLaughlin W., Blake W., 1979. "Rock preconditioning to Prevent Rockbursts - Report on a Field Demonstration", USBM RI8318.
Koopmans, R., Hughes, R.W., 1984, "An Automat~d and Comp~terized Dilatometer System to Measure Deformation ModuluB of Rock", Western States Mining Expo, Nevada.
Koopmans R., Hughes R., 1985. "Determination of Near Field Excavation Disturbance in Crystalline Rock", 9th Int. Symp. on Scientific Disturbance in Crystalline Rock, Sweden.
Kutter H., Fairhurst C., 1971. "On the Fracture Proces8 in Blasting", Int. J. Rock Mech. Min. Sci., Vol. 8. pp. 181-201.
, 197 ,
c
J
c:
,.
Kutter H., Fairhurst C., 1968. "The Role of Stress Wave and Gas Pressure in Presplitting". cProc . 9 Rock Mech. Sym., Colarado. AIME, New York.~ pp. 265-284.
Ladanyi, B., 1975. "Quasi Static Expansion of a Cylindrical Cavity in Rock", Proc. 3rd Sym., Eng. Application of Sol id Mech., Toronto. pp. 219-240.
Ladanyi, B., Gill, D.E., 1981. "Determination of Creep Parameters of Rock Salt by Means of a Borehole Dilatometer", Proc. 1st Conf. on Mech. Behavior of Salt, Penn. State Univ., Trans Tech. Pub., Germany. pp. 267-273. )
Lama R.D., 1972. "The Use of ifestressing Techniques in the Maintenance of Roadways Under High Pressure". Proc. Int. Sym on Underground Openings, Lucerne. pp. 177-181.
Lama R.D., 1974. "The Uniaxial Compressive Strength of Jointed Rocks". Prof. L. MuLLer Festschrift Inst. Soil Mech. and Rock Mech, Univ. Karlsruhe, Karlsruhe.
Lama R.D., Vutukyri V.S., 1978. "Hanbook on Mechanical Properties of Rocks: vol. ii, iii, iv". Trans Tech Pub., Germany.
Le Bel G.R., Quesnel W.J.F., Glover W.,1987."An Analysis of Rockburst Evénts During Sinking of Macassa 13 Shaft". Proc. CIM AGM, Toronto.
Leighton F.,~984. "Microseismic ~onitoring and Warning of RQckbursts". Proc. lst Int Congo on Rockbursts and Seismicity in Mines, SAIMM, Johannesburg. pp.287-296.
Leighton F., 19B2a. "Dynamic Methods of rock Structure Analysis". Underground Mining Methods HanQ~, chap. 7, W.A. Hustrulid (ed.), Society of Mi~ing Engineers, AIME, N.Y. pp . 1523 -15 3 4 .
Leighton F.,1982b. "Case History of a Major Rockburst" USBM RI Love A.E.H., 1927. "A Treatise on the Mathematical Theory of
Elasticity". Cambridge, Cambridge University Press. Makuch A., Neumann M., Hedley D., Blake W.i 1987. "Practical
Applications of Pillar Destressing at Campbell Red Lake Mine", Presented at Bth Underground Oper. Conf., CIM.
Makuch A., 1986. "Field Stress M~asurements". InternaI Report Campbell Red Lake Mine
McGarr A., 1984. "Sorne Applications of Seismic Source Mechanism Studies to Assessing Underground Hazard". Proc. Rockburst and Seismicity, SAIMM, Johannesburg. pp. 199-208.
Mohanty B' f 1988. Personal Communication. Mohanty B' f 1982. "Characteristic Crack Patterns Close to an
Exploding Charge". 14th Cdn. Rock Mech. Sym., CIM Special Vol. 30. pp. 32-36.
Moruzi G., pasieka A., 1964. "Evaluation of a Blasting Technique for Destressing", Proc. 6th U.S. Symp. Rock Mech., Univ. Missouri. pp. 185-204.
Morrison R., 1961. "A Report on Rockburst Potentiality at Dickenson Mines Ltd. and Campbell Red Lake Mines Ltd.", Consultants Report.
Neumann M., Hedley O., Blake W., Makuch A., 1987. "An Instrumented Destressing Blast in the 1604 EW Stope at Campbell Red Lake Mines", CANMET, Ontario/lndustry/Canada Rockburst Research Project.
198
o
,
Neumann M., Makuch T.,l985. "Case Study of Microseiamic Monitoring of the 'F-2' Zone at Campbell Red Lake Mines Limited". Proc. Microseismic Monitoring in Canadian Mines, CANMET., MRP/MRL 85-23.
Neumann M.,1986. Personal communication. Obert L., Duvall W. 1 1957. "Microseismic Methods of Determinihg
the Stability of Underground Openings". USBM Bul. 573. 0bert L., f)uvall W., 1945. "Microseismic Méthod of Predict ing' Rock
Failure in Underground Mining: Part I~~ Laboratory Experiments". USBM. RI3803. ~
Oliver P., Wiles T., MacDonald ~., 0'Donne1l D., 1987. "Rockburst Control Measures At INCO's Creighton Mine. Presented at CIH AGM, Toronto.
Olsen J.J., Willard R.J., Fogelson D.E., Hjelmstad K.E., 1973. "Rock Damage from Small Charge Blasting in Granite", USBl1 R1 7751.
Ontario Government, ,1986. "Improving Ground Stability and Mine Rescue: The Report of the Provincial Enquirey into Ground ,~ Control and Emergency Prepardness In Ontario Mines". Ontario Government Puplication.
Quesnel W.J.F., Hong R., 1986. "The Use Of Rock Mass Conditioning: Destressing For The Safe Recovery Of Rockburst Prone'Pillars - Macassa Mine". Ground Control Seminar,
'Hailibury School Of Mines. Quesnel W.J.F., 1987. Personel communication. Rigg D. M., 1980. "Relationships Between Structure and Gold
~ Mineralization in Campbell Red Lake and Dickenson Mines, Red take District, Ontario". M.Sc. Thesis, Queen's University.
Robson W., 1946. "Rockburst Incidence, Research and Control Measures at Lake Shore Mines Ltd.". Trans. CIM., Vol. XLIX. pp. 347-375.
Rocha M., Silveira A., Peres-Rodrigues F., Silvero A., Ferreira A., 1970, "Characterization of the Deformability of Rock Masses by Dilatometer Tests", Proc. 2nd ISRM Cong., Belgrade. pp. 2-32.
Roctest, 1986. "Probex-1 Dilatometer Instruction Manual", Roctest Ltd., Montreal.
Roux A., Leeman E., Denkhause H., 1957. "Destressing: a Means Of Ameliorating Rockburst Conditions. Part 1: The Concept of Destressing and the Results Obtained From Its Application". Journal South African Inst. Min. Met, Vol. 56. pp. 241-267.
Salomon M., 1984. "Energy Considerations in Rock Mechanics: Fundamental Results". J. SAIMM, vol 84, No. 8. pp. 233-246.
Salomom M.,1983. "Rockburst Hazard and the Fight for its Alleviation in South African Gold Mines". Proc. Rockburst Predit ion and Control, IMM, London. pp. 11-36.
Scoble M.J., Cùllen M., Makuch A., 1987, "Experimental Studies of Factors Relating to Destress Blasting", Proc. 28th U.S. Symp. Rock Mech., Tuscon. AA Balkema, Bos.ton. pp. 901-908.
Scoble M.J., Cullen M., Carter M., Barclay R.,1988. "In Situ Deformability and Fracturation Studies Using The Probex-l Borehole Dilatometer". Proc. Conf. on Applied Rock Eng., New Castle. IMM, London. pp. 203-212.
199
c -1
8 Siskind D.E., Fumanti R.R., 1974. "Blast-Produced Fractures in
Lithonia Granite", USBM R1 7901. Siskind D.E., Steckley R.C., OIson J.J., 1973. "Fracturing in the
Zone Around a Blasthole, White Pine, Mich." USBM. RI 7753. Slade A., 1968. "Sorne Practical Aspects of Rock Mechanics in
Falconbridge's Sudbury Operations". CIM Bul., Vol. 61, No. 678. pp. 1161-1165.
Timoshenko S.P., Goodier J.N., 1970. "Theory of Elasticity". Wiley, New York.
Vutukuri V.S., Lama R.D., Saluja S.S.,' 1974."Mechanical Properties of Rocks: vol. i". Trans Tech Pub., Germany.
Walsh J.B.,1965a. "The Effects on the Unixial Elastic . Compresssion of Rqcks". Journal Geophys. Res., Vol 70, No. 2. pp. 399-411.
W~lsh J.B.,1965b. "The Effect of Cracks in Rocks on Poisson's Ratio". Journal Geophys. Res., Vol 70, No. 2. pp. 381-389.
Walsh J. B., 1965c. "The Effect of Cracks on the Compressibility of Rock" . Journal Geophys. Res., Vol 70, No. 20. pp. 5249-5257.
j Willan J., Scoble M., Pakalnis V,'
1985. "Destressing Practice in Rockburst Prone Ground", Proc. 4th Conf. Ground Control in Mining, W. Virginia Univ.
Young P.,1986. "Seismic Propagation in Rock Masses: Applications for Acoustic Ernissions/Microseismic Monitoring". 4th Conf. Acoustic Emissions/Microseismic Activity in Geologic Structures and Materials. Trans Tech Pub., Germany.
/
...
i
200 ,