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8/3/2019 Newsletter 33 Web
1/27
The Captain's pit in Malmberget. Photo courtesy of LKAB
A U S T R A L I A N C E N T R E F O R G E O M E C H A N I C S V o l u m e N o . 3 3 D e c e m b e r 2 0 0 9
NEWSLETTER
The views expressed in this newsletter are those ofthe authors and may not necessarily reflect thoseof the Australian Centre for Geomechanics.
Continued page 2
Sublevel caving past and
future
IN THIS EDITION Sublevel caving past and future, Page 1 In-pit risks, Page 7 Mine closure planning, Page 11 Mining-induced seismicity, Page 15 Tailings disposal, Page 17 Mine tailing solutions, Page 20 Increasing value of paste, Page 21 ACG event schedule, Page 24
by William Hustrulid, University of Utah; and theColorado School of Mines, USA, and Rudolph Kvapil, USA
www.minewaste2010.com
29 September 1 October 2010,
Perth, Western Australia
Abstracts due 1 March 2010
IntroductionThe sublevel caving technique according to
early mining books (Peele, 1918) evolved in
the U.S. from top slicing. It was a logical next
step in the mine geometry scale-up process.
Block caving, in turn, was the logical scale-up
from sublevel caving.
Janelid (1972) indicates, In the rst
application of sublevel caving, the ore was not
drilled and blasted completely between twosublevels, but certain parts were broken by
induced caving (hence the name sublevel caving).
As the method is applied today, the whole
quantity of ore between the different sublevels
is broken (or at least should be) using controlled
drilling and blasting. If this is done in a proper
and rational way, there are good possibilities
of developing a mining method which can be
applied, technically as well as economically, on
any orebody of suitable size, location and rock
mechanical properties .
In spite of some searching, the modern
origins of todays version could not be
clearly identied. Possibly it was developed
in the iron mines of Sweden. Janelid (1972)
indicates, For a long time, sublevel caving wasthe predominant mining method at Grngesberg.
During the last ten years (since about 1960),
however, block caving has given 70% of the
production.
In 1960, the sublevel caving technique
was being used by 19 Swedish mines with a
Mine Waste 2010 will tackle the full
range of issues that constitute risks
in the management of mining wastes,
particularly tailings and waste risk.
This forum will encourage debate
amongst practitioners, researchers and
regulators about the key shortcomings
in industrys current understanding
of the performance of mining waste
storage facilities and associated risks
faced by owners and operators of these
facilities.
First International Seminar onthe Reduction of Risk in the
Management of Tailingsand Mine Waste
8/3/2019 Newsletter 33 Web
2/272 Australian Centre for Geomechanics December 2009 Newsletter
Continued from page 1
Copyright 2009. Australian Centre for Geomechanics (ACG), The University of Western Australia (UWA). All rights reserved. No part of this newsletter
may be reproduced, stored or transmitted in any form without the prior written permission of the Australian Centre for Geomechanics, The University of
Western Australia.
The information contained in this newsletter is for general educational and informative purposes only. Except to the extent required by law, UWA and the
ACG make no representations or warranties express or implied as to the accuracy, reliability or completeness of the information contained therein.
To the extent permitted by law, UWA and the ACG exclude all liability for loss or damage of any kind at all (including indirect or consequential loss ordamage) arising from the information in this newsletter or use of such information. You acknowledge that the information provided in this newsletter is
to assist you with undertaking your own enquiries and analyses and that you should seek independent professional advice before acting in reliance on the
information contained therein.
the economic benets which can be achieved
through the development of the correct
method are extraordinarily large.
In Czechoslovakia in 1950, Rudolf
Kvapil was given the task of determiningthe causes of problems in bins and silos
and, based on this new understanding,
to develop ways of improving their
performance. It was evident to him that it
would rst be necessary to determine the
basic gravity ow principles for granular
and coarse materials since they must
be completely different from principles
describing the ow of liquids which were
then available for use. He decided that
the only realistic way to proceed was
to construct and test a large number of
models and to make in situ observations.
Many of these models and the knowledgegained are described in his recent book
(Kvapil, 2004). In 1965, Kvapil joined Janelid
at KTH and began applying the gravity ow
principles gained in the study of bins and
silos to sublevel caving.
Figure 2 shows the application of this
type of model to a sublevel cave design. In
this particular case, the sublevel spacing is
12.5 m, the drift dimension is 5 x 3.5 m,
the sublevel drift spacing is 12 m and the
burden is 2 m. These closely resemble the
sublevel dimensions used by the Kiruna
Mine in the early 1980s. It is interestingto note that the design is based on a
drawbody width (WT) to drawpoint width
(WD) ratio of 1.7.
Figure 2 Application of gravity flow principles tosublevel caving design (Kvapil,1982, 1992)
Over the past few years, the scale of
sublevel caving has increased markedly with
LKAB being a leader in this regard. Figure
3 provides a comparison of the sublevel
caving mining geometries appropriate
for the years 1963, 1983 and 2003 at
the Kiruna Mine. Some of the importantparameters are tabulated in Table 1.
Figure 3 The sublevel caving geometry at the KirunaMine at three different points in time (Marklund andHustrulid, 1995)
At the Kiruna Mine today the sublevelspacing is 28.5 m. In certain sectors of
LKABs Malmberget Mine, the sublevel
spacing is as high as 30 m.
Table 1 Summary of some important
design parameters (Marklund and
Hustrulid, 1995)
Today, with the continuing push to
increase mining scale, a fundamental
question is whether the gravity ow
principles which served as the design
basis for the small-scale sublevel caving
mine designs of the past can be applied
at much larger scales or whether some
other approach is required. This article willprovide some thinking in that regard.
total yearly production of about 9.5 Mton
(Ohlsson, 1961). Figure 1 is a sketch of
the method as practiced at LKABs Kiruna
Mine at about that point in time.
Figure 1 Composite section view of the sublevelcaving mine at Kiruna in 1957
The scale was small, certainly by todays
standards, with a sublevel spacing of 9 m, a
drift size of 5 x 3.5 m, and a sublevel drift
spacing of 10 m centre-to-centre.
As Janelid (1961, 1972) pointed out,
Sublevel caving is in many respects simple. It
can be used in orebodies with very different
properties and it is easy to mechanize.
However, from other points of view such as
recovery, dilution and similar, the method is
unfavorable. The designs which are used andthe measures which can be taken to eliminate
the disadvantages are poorly understood. In
the end of the 1950s, model tests regarding
gravity ow in material resembling broken
rock were started at the Division of Mining,
the Royal Institute of Technology (KTH) in
Stockholm. The purpose was to study how the
geometrical design of various parameters in
sublevel caving are inuenced by the motion
which is induced in the material when ore
is loaded in a sublevel drift. Some of these
model tests were performed as a part of
senior theses and others by assistants and
research engineers. Model tests and extensiveliterature studies on sublevel caving have
also been carried out in Kiruna together with
conducting practical tests underground. The
results achieved have been so encouraging that
continued research work is well justied since
Year
Parameter 1963 1983 2003
Drift width (m) 5 5 7
Drift height (m) 3.5 4 5
Sublevel height (m) 9 12 27
Sublevel drift spacing
(m) 10 11 25
Blasthole diameter(mm) 45 57-76 115
Burden (m) 1.6 1.8 3
Holes/ring 9 9 10
Tons/ring (t) 660 1080 9300Tons/metre of drift
(t/m) 400 600 3100
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CavingMine marker studies
It is one thing to study ow principles
in a laboratory setting and quite another
to show that they apply in the reality ofa mine setting. One way of doing this is
through marker studies. Figure 4 shows
some results from the rst marker studies
conducted as part of the overall KTH
sublevel caving research programme
conducted at the Grngesberg iron mine in
central Sweden in the early 1970s.
Figure 4 Results of the Grngesberg marker tests(Janelid, 1972)
Some of the relevant parameters are
summarised in Table 2.
Table 2 Design parameters at Grngesberg
From Figure 4, it appears that the ow
width is of the order of 5 m. Since the
drift width is 3.5 m, the ow width to
drift width ratio is 1.43. Due to the roof
curvature, the effective extraction width
is somewhat less and the ratio would be
corresponding slightly larger.
It took quite a long time for the
next group of mine marker tests to be
performed. As noted by Quinteiro et al.
(2001), The sublevel caving layout used at
Kiruna has reached dimensions that are far
beyond those that formed the basis for thedevelopment of the early design guidelines.
Thus, there was a need to verify the gravity
ow pattern for this very large sublevel caving
area. It was decided to install markers in the
fans so one could estimate the ellipsoid of
extraction.
Figure 5 shows the fan geometry and
Table 3 summarises some of the importantparameters.
Figure 5 Fan geometry for the Kiruna sublevel cave
Table 3 Summary of some important
factors concerning the Kiruna marker tests
Figure 6 shows the results of the
recovered markers expressed as a
percentage of the total number of markers
installed at each particular location.
Figure 6 Percentage of the recovered markers at a
particular position
It can be seen that only a very small
number of markers were recovered from
the sides of the fan indicating that the ore
ow was small. On the other hand, a large
number of markers were recovered from
the central part of the fan indicating that
the predominant ore ow pattern was inthe center. This type of ow behavior will
result in early dilution. Figure 7 shows the
results in Figure 6 in the form of a contour
plot.
Figure 7 Contour plots showing the percentrecoveries at the different marker positions
Recently, comprehensive marker studies
have been carried out at the Perseverance
and Ridgeway sublevel caving mines inAustralia. At the Perseverance Mine, the
overall ow pattern as demonstrated using
the markers is shown in Figure 8. Some of
the important parameters are presented in
Table 4.
Figure 8 Section showing the rings with the drawpattern superimposed. Perseverance Mine
Parameter Value
Sublevel drift spacing (m) 7
Sublevel spacing (m) 13
Hole diameter (mm) 41
Burden (m) 1.5
Sublevel drift width (m) 3.0 slashed to 3.5
Sublevel drift height (m) 3
Front inclination (degrees) 90
Parameter Value
Sublevel drift spacing (m) 25
Sublevel spacing (m) 27
Hole diameter (mm) 114
Burden (m) 3
Sublevel drift width (m) 7
Sublevel drift height (m) 5
Front inclination (degrees) 80
It is one thing to study ow
principles in a laboratory
setting and quite another toshow that they apply in the
reality of a mine setting.
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Table 4 Summary of some important
factors concerning the Perseverance
marker tests
Table 5 summarises some of the
important parameters concerning the
Ridgeway marker tests.
Table 5 Summary of design parametersfrom the Ridgeway Mine
In reviewing the results of the marker
tests from the Grngesberg, Kiruna,Perseverance and Ridgeway mines, it is
interesting to note that they all basically
reveal a type of silo ow such as shown
in Figure 9 even if the drilling pattern
extends far outside of the silo.
Figure 9 Silo type of flow pattern. Kvapil (1955),Janelid and Kapil (1965)
The average primary ow width/drift
width ratios (Wf/Wd) for the four casesare summarised in Table 6.
Table 6 A comparison of the marker ow
The Wf/Wd ratio of 1.4 1.7 seems
to apply for small scale sublevel caving
geometries as well as very large scale.
These results are in agreement with
the early sublevel caving geometry
recommended by Kvapil (see Figure 2)
which used 1.7.In retrospect, there are three reasons
why this is a very logical result:
1. The middle holes of the ring are red
rst and can make rst use of the
swell volume offered by the underlying
sublevel drift.
2. The central holes are drilled subvertical,
fairly parallel, and relatively close to one
another. The result is a relatively high
and uniform specic charge compared to
the other holes in the round. Thus, one
would expect the best, most uniform
fragmentation.3. The ore material in the central part of
the round can make the best use of the
effect of gravity in directing it to the
drawpoint.
As indicated earlier, small-scale physical
model test results have historically played
a very important role in the dimensioning
of sublevel caves. In the construction of
these models, the sand or other material is
simply poured into the forms. As such, the
properties are uniform and the mobilities
are the same independent of position
within the model. In a sublevel cave, this is
not the case. All of the material in the fanis drilled and blasted. Because of the fan
geometry, the amount of explosive/unit
volume and hence the fragmentation varies
throughout the fan. The ore material in the
centre part of the fan and the lower part
of the fan has a much higher specic charge
than that at the boundaries of the ring.
Furthermore, the cave which lies in front
of the blasted slice is an eclectic mixture of
waste rock and ore remnants. Its mobility
varies with location and with time (it
changes with the extraction geometry).
Finally, most rock materials upon beingblasted would like to bulk (swell) of the
order of 50%. In sublevel caving, it is the
sublevel drift located at the bottom end
of the fan which is the primary provider
of swell space for the ore in the ring. As
shown in Table 7, the available free swell is
highly mining scale dependent.
Table 7 Available free swell for the
different LKAB designs
As the scale has increased over the
years in the quest to reduce the specic
development, the available free swell has
correspondingly decreased. With the
current LKAB design it is only about 5%.
Since it is located at the bottom of thefan, the ore in the near vicinity of the drift
has a much greater access to this volume
and the chance to bulk. The ore at the
extremities of the fan, on the other hand,
has little chance to bulk and its mobility is
very low. Based on material mobility alone,
one would expect signicant differences
in the mechanics of ow between the
sand models and reality, particularly as
the sublevel scale is increased. Hence,
the marker test results have very high
signicance.
Sublevel cave layout rules basedupon marker test input
Based upon the results of the four
marker tests, it appears that the Wf can be
expressed as a constant times the width of
the Wd. As a rst approximation,
Wf= (1.4 1.7) Wd (1)
Some preliminary design rules for initial
planning are summarised below:
Sublevel drift size (width (Wd) and
height (Hd): determined based on
equipment. Sublevel interval (HS): the theoretical
maximum value is based on the ability
Parameter Value
Sublevel drift spacing (m) 14.5
Sublevel spacing (m) 25
Hole diameter (mm) 102
Burden (m) 3
Sublevel drift width (m) 5.1
Sublevel drift height (m) 4.8Front inclination
(degrees) 75
Parameter Value
Sublevel drift spacing (m) 14
Sublevel spacing (m) 30
Hole diameter (mm) 102
Burden (m) 2.6
Sublevel drift width (m) 6
Sublevel drift height (m) 4.7Front inclination
(degrees) 80
Mine
Drift width
(Wd)
Level
interval
Flow width
(Wf) Wf/Wd
(m) (m) (m)Grngesberg 3.5 13 4.9 1.4
Kiruna 7 27 10.3* 1.5
Perseverance 5 25 7.1 1.4
Ridgeway 5.9 25 - 30 10.0 1.7
* Arbitrarily taken as the 30% contour
Design "Free" Swell
1963 24.0
1983 17.9
2003 5.5
patterns
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Cavingto drill long, straight holes. This, in turn,
is based on the hole diameter (D). The
actual limit is based on recovery and
dilution considerations which are due to
managing ore/waste pulsation. Hole diameter (D): based on the
available drilling equipment and the
ability to charge long holes.
Spacing of the sublevel drifts (Sd):
Sd = (2.4 2.7) Wd (2)
Ring spacing (burden (B)): based
upon the damage radius (Rd) concept
discussed by Hustrulid and Johnson
(2008)
B = 2 Rd (3)
Where:
(4)
Rd
= damage radius (m)
rh
= hole radius (m)
PeExp
= explosion pressure for the
explosive
Pe ANFO
= explosion pressure for ANFO =
1600 MPa
rock = rock density (g/cm3
)2.65 = density of typical rock (g/cm3)
Hole toe spacing (ST): based upon the
burden
ST
= 1.3 B
Spacing for parallel holes (SP): based
upon the burden
SP = B (5)
Front inclination: 7080 degrees
(forward)
If it is assumed that:
D = 115 mm
Drift dimensions: 7 m wide by 5 m highExplosive: emulsion (P
e Exp= 3900 MPa)
Rock density = 4.6 g/cm3
Sublevel interval: 25 m based on drilling
ability and control of pulsation
One nds that the remaining dimensions
are:
Sublevel drift spacing: 1719 m
Burden: 2.7 m
Toe spacing (fanned): 3.5 m
Toe spacing (parallel): 3 m
Front inclination: 80o selected
It is noted that the new sublevel drift
spacing rule has very limited basis and mustbe carefully complemented with further
testing.
Implications for future sublevelcaving designs
The results of the marker studies would
suggest that modications in some of thecurrent, very large scale sublevel caving
designs should be considered. Assuming
that the drift width is not changed, the
results suggest that the sublevel drift
spacing should be reduced. Presuming
that there is no change in the sublevel
height, this means that the overall mining
scale would decrease and the specic
development would increase. One way of
maintaining the current scale is to increase
the width of the sublevel drift. Figure 10
shows one possibility.
Figure 10 Silo design with super-scale extractiondrifts, patterned after Kvapil (1992)
This has advantages with respect to the
silo shape and the parallel hole drilling.
However, one must be concerned with
geomechanics issues (drift and brow
stability). Furthermore, the draw must be
well controlled over the entire face.
If one wants to preserve the specic
development ratios in place today, one
would need to increase the sublevel
height. However, this has problems with
hole deviation, maintenance of long holes,
charging of very long holes, and dealing
with ore/waste pulsation over a much
longer draw duration. This seems like avery difcult alternative to achieve on a
day-to-day basis. On this basis, it would
seem that in the future mining companies
will be looking toward smaller scale designs
than today and not larger. The current very
large-scale designs may actually be too
large-scale.
Front caving implications
This article has only dealt with standard
sublevel caving. There are a number of
variants, however. Front caving is a variety
of the sublevel caving technique which isquite often used. It is, for example, a very
interesting technique for the creation of
the undercut required in block and panel
caving. However, it is very important that
the undercut be completely formed. The
marker studies would indicate that the ow
stream is much narrower than previously
thought. If rock mass ow does not occurover the full drilled width, the remaining
portions could form remnants and
transmit loads to the production level with
catastrophic consequences. This means that
current undercut designs based upon front
caving will have to be re-evaluated.
Future possibilities to maintain/increase scale
There are two possibilities, at least, to
try and maintain or possibly even grow
the scales used today. One possibility
deals with using more of the sublevel drift
for swell than just that taken by the orefalling down. This involves changing the
blasting pattern and initiation sequence
so that the ore at the lower part of the
ring is propelled far out into the drift. A
second possibility which also involves a
change in the blasting is to use the available
swell space more effectively. This means
permitting the ore in the lower part of the
ring to only swell 20%, rather than 50%.
This would thereby increase by a factor of
2.5 the amount of ore in the ring which has
a chance to swell. Accomplishing both of
these possibilities should be well within thecapabilities of electronic detonators with
very precise timing.
A problem with todays typical ring
drilling design is that the hole spacing
changes from very small near the drift to
large at the hole ends. The parallel hole
design used in the silo design avoids this
problem. Without a major change in drift
width, one is conned to a rather narrow
pattern. Figure 11 shows one possible
futuristic design involving special drilling
technology and the blasting innovations
which better use the available free swell
space.The design presents an opportunity
to achieve improved fragmentation, an
increase in ore mobility, and a more
uniform distribution of ore mobility over a
much wider front. An understanding of how
the ore actually ows in sublevel caving will
lead to better designs. The marker studies
are an important step along that path.
rockANFOe
Expe
hdP
PrR
65.220/ =
The results of the markerstudies would suggest that
modications in some of
the current, very large scalesublevel caving designsshould be considered.
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Second International
Symposium on Block andSublevel Caving
2022 April 2010,
Novotel Langley Hotel,
Perth, Australia
The growing popularity of cavingmethods around the world is largelydue to the very low production costand the intrinsic safety associatedwith this mining approach. Morethan 50 technical papers areexpected to be presented at this
three day event.
www.caving2010.com
2010CAVING
5thInternational Seminar onDeep and High Stress Mining
2010
68 October | Santiago - CHILE
Ponticia Universidad Catlica
de Chile, in collaborationwith the Australian Centre forGeomechanics, the Universityof Toronto, and the University ofWitwatersrand, is organising anInternational Seminar on Deepand High Stress Mining.
As the mining industryfaces new challenges toextract mineral resources atincreasing depths, the DeepMining International Seminarseries provides a forum for
the industry, academicsand researchers to shareinformation, experience andideas on deep and high stressmining.
For more details [email protected] or visit http://web.ing.puc.cl/~deepmining2010/
Collaborating Organisations
William HustrulidUniversity of Utah; andthe Colorado School ofMines, USA
Future studiesIn closing, the authors believe that
it is time to seriously revisit the
recommendation made by Janelid (1961)
nearly 50 years ago with regard to small-
scale sublevel caving,The results achieved
have been so encouraging that continued
research work is well justied since the
economic benets which can be achieved
through the development of the correct
method are extraordinarily large.
In spite of their obvious value, eld
studies are few and far between in the
mining business. In addition, if conducted,it is very difcult for others to access the
results and perhaps gain and offer new
insights. This must change if the mining
business is to meet the technical, economic
and safety challenges the future has to offer.
There is a real danger that todays
sublevel caving designs are far from
optimum due to a poor understanding
of the fundamental processes involved.
In the past, the application of sublevel
caving has primarily been to iron ore,
particularly magnetite, which because of its
very forgiving magnetic property, permits
easy and inexpensive separation from thewaste. The same is not true with other
minerals, for example copper porphyry and
gold ores. For these, it is very expensive
to separate ore and waste. It would
appear that prior to fully committing to
any sublevel caving design, a pilot project
should be run with a carefully planned and
executed program of data collection. One
very important piece of information to
be extracted is the draw width. It is also
very important to develop the required
draw control techniques to be applied in
the mine. Ore/waste pulsation, which isinherent in very high draw designs, makes
practical draw control very difcult. Visual
viewing of the cave front is not enough.
AcknowledgementThis edited article is from the paper
entitled, Sublevel caving - past and
present featured in the proceedings of
the 5th International Conference and
Exhibition on Mass Mining, Lulea, Sweden,
911 June 2008.
Figure 11 New possibilities for large-scale sublevel caving
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Introduction
Risk, risk assessment and risk analysis
have a number of meanings across a range
of disciplines. At the most fundamental,
risk is simply a combination of uncertainty
in an outcome and consequences for that
outcome. Risk analysis or risk assessmentis the process of identifying, quantifying,
and communicating those uncertainties
and outcomes. In geological engineering,
risk has traditionally been tied to the
calculation of a factor of safety of a slope,
or potential failure geometry, and has
historically been a qualitative assessment
of a calculated value. Advances in the
computational power of stability analysis
software programs have set the stage for
more quantitative assessments. Depending
on the scale of the slope under evaluation,
and given the variation inherent in earthmaterials in general, almost every input
can be considered to vary over a range of
potential values.
As such, risk assessment in geological
engineering often considers both aleatory
uncertainty - the variability inherent to
natural materials, and epistemic uncertainty
- the variability related to the ability to
model a phenomenon. It is uncommon,
however, that risk assessment considers a
temporal element, i.e. how the inputs, and
therefore the associated risk, change with
time. To an extent this is to be expected
as many inputs do not signicantly changeover the course of a project life. However,
elements such as pore pressure, the surface
topography of an excavation, the weight
distribution on a potential failure plane,
the probability of a seismic event and the
properties of low strength materials can
all change to a magnitude that materially
affects the outcome of a risk analysis. No
attempt has been made in this assessment
to look at equipment or personnel
temporal exposure.
To evaluate the effect of the aleatory,
epistemic and temporal variation, researchwas conducted at the Rio Tinto Minerals
Boron Operations open pit mine near
Boron, California. The purpose of this
The changing prole of risk associated
with in-pit placement of waste
orebody that was deposited as an evaporate
by Raymond Yost, Rio Tinto Minerals Boron Operations, USA
article is to discuss the background to that
work, the nature of the risk analysis and
assessment, and to present preliminary
results.
Background and site characterisation
The Boron open pit mine is located
near the town of Boron, California in theMojave Desert Geologic Province. The
mining operation extracts borates from
a lenticular orebody that was deposited
as an evaporite and is encased in layers
of low permeability claystone. The clay
and borate sequence is bounded on the
bottom by a layer of basalt, which is in
turn underlain by feldspar-rich sandstone
(arkose) with interbeds of clayey sand (the
Tropico Formation). Poorly to moderately
consolidated and cemented arkose covers
the borate and clay sequence. An intrusive
body, composed primarily of quartzmonzonite, bounds the deposit to the
south.
The sequence of Tropico-basalt-
evaporites-sediments has been tilted and
dips moderately; 5 to 15 to the south.
Faulting has offset the orebody into three
primary components and a number of
sub-blocks.
The open pit operation was initiated in
the late 1950s in the northwestern portion
of the deposit where the borate layer was
generally closest to the surface. Over the
past 60 years, the pit has expanded to the
south and east and has deepened as thehigher elevation ores have been mined out.
Slope failures that have occurred during
open pit mining operations typically form
due to a combination of pore pressure,
high-angle faults (which act as a back plane)
and low-strength beds of clayey sand or
claystone. All of the open pit slopes are
designed in recognition of these variables.
The design of the north wall, however,
is also governed by the orientation of
the orebody. As offset on most faults is
relatively minimal, the overall slope of
the wall generally follows the overallorientation of the orebody.
The overall slope angle of the wall,
in conjunction with the strength of the
foundation material (basalt), generally
results in factors of safety well in excess
of industry required limits. Furthermore,
the mineralised zone at the site is conned
to a single geologic unit. Extraction of
the borate layer represents complete
extraction of the resource, so dumping
over mined out areas does not presentany risk of covering potentially economic
mineralised zones. The north slope of the
pit was therefore an attractive option
for overburden disposal given that it was
stable, composed of a higher strength unit,
and close to active mining operations. A
risk assessment was conducted prior to
the large-scale placement of overburden on
the slope.
Structure of the risk assessment and
input variables
Mining in the most general sense,balances two basic elements benets
realised against the potential for loss. In
this case, they have been incorporated
into the risk assessment. Benets are
realised if the ground and overburden
dump remain stable throughout the project
life and costs are incurred if they do not.
Evaluating risk in this case is therefore a
matter of determining the potential for
slope instability along with the values
of the benets and costs. Stability is a
function of the geology, the potential
for a seismic event, the pore pressure,
the size of the dumped volume and theslope conguration. While some of these
variables remain constant over the project
life, most of them change to a large enough
degree that they affect the probability of a
slope failure. A thorough risk assessment
therefore requires an evaluation of
conditions through the full time line of the
project.
The risk assessment was structured to
evaluate the potential for slope failure.
The risk through time was quantied via
a series of steps to establish a probability
of failure, determine the magnitude ofpotential negative outcomes and model the
expected values. Specic tasks included:
Open pit
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1) Estimating the probability of an outcome
(a slope failure) through the use of
limit equilibrium analysis and statistical
sampling of analysis inputs.
2) Estimating the likely extent of negativeresults (failure clean-up) through the
use of numerical and empirical methods
to develop a model of post-failure
topography.
3) Estimating the likely extent of positive
results (savings associated with dumping
near the area of extraction as opposed
to ex-pit dumps) through an evaluation
of equivalent tonne miles (ETM).
4) Using the probability of an outcome
and the estimated costs and benets to
establish expected costs and benets
with time.
5) Adjusting the timing of benets and costs(benets are expected to be realised
early while costs are expected to be
realised later) with a discount rate.
6) Estimating a net expected sum of
benets at distinct points.
Once these values were estimated,
the risk was determined as the net sum
of expected benets and costs. A value
greater than zero implied that the outcome
had a positive expected economic value,
while a net sum of less than or equal
to one implies that the outcome had anegative expected economic value and a
negative risk. The evaluation was repeated
at appropriate time increments for a range
of in-pit dump volumes to determine if, and
how much, waste could be economically
placed in the pit.
Results
To illustrate the interplay of the
various inputs to the risk assessment,
the start and end points of one of the
analyses are presented in Figure 1, from
the limit equilibrium analysis through
empirical modelling, to the nal economicassessment for a 30 million t in-pit waste
dump.
Risk in the most general sense,balances two basic elements
benets realised against the
potential for loss.
Figure 1 Pit topography at year 2010
Probability of failure 5.0% (with seismic load).
Probability of failure 0.55% (without seismic load).
Failure volume (in section) 28,350 m 3 (with seismic load).
Failure volume (in section) 28,700 m 3 (without seismic load).
Probability of seismic event 5.0%.
At this beginning stage, ore (blue and green units) is close to the toe of potential
failure and subject to burial should failure occur. Failure volume is relatively high, but
the probability of failure is relatively low. The probability of a seismic event occurring is
relatively low.
Figure 2 Pit topography for ultimate pit
Probability of failure 81.20% (with seismic load).
Probability of failure 41.60% (without seismic load).
Failure volume (in section) 39,500 m 3 (with seismic load).
Failure volume (in section) 39,250 m3
(without seismic load).Probability of seismic event 70.0%.
At this nal stage, failure volume increased by approximately 40%, but the probability of
failure increased, on average, to approximately 60%. The potential for a seismic event has
increased as well, but the ore zone is farther away from the toe of slope and is less likely to
be covered by a slope failure.
Modelling post-failure runout
A combination of numerical modelling and empirical evaluation was used to develop
potential post-failure topography. Post-failure proles were developed for all sections with a
probability of failure greater than 0.01% regardless of the factor of safety. The conguration
of the runout was based on an assessment of historical slope failures at the site. At Boron
this was the angle of repose of the failed material relative to the dip angle of the underlying
failure plane, and adjusted for the geometry of the runout area.
Figure 3 Topography for failure at 2010 for 30 million tonne dump
The ratio of the clean-up area to the post-failure area is 18.5%. The runout was contained
to some extent by the concave geometry of runout area resulting in a low overall angle of
repose.
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Figure 4 Topography for failure at ultimate pit for 30 million tonne dump
The ratio of the clean-up area to the post-failure area is 22.5%. The removal of material
below the toe of failure has allowed considerable runout. The overall angle of repose has
increased.
Benets and costs
Assessing benets and costs began with establishing values for dumping a unit of waste in
the pit and for cleaning up a unit of failure debris from the pit. The value of dumping tonnes
in the pit is a function of reducing both horizontal and vertical haul distances. Reducing
the haul distance generally means that additional truck hours are available. These truck
hours are either used to haul additional waste, or, if enough truck hours are offset by the
short hauls, a truck(s) could be parked. The difference in either case is reected by overalllower haulage costs. The problem lies in translating these lower overall costs into what the
specic unit cost difference is for dumping a portion of the waste in the pit versus hauling
all waste outside of the pit.
To accomplish this, it was necessary to evaluate haul costs with a unit that accounted
for both the difference in horizontal and vertical travel distances associated with hauling
to a site outside of the pit, versus hauling to a site inside the pit. The value used was the
ETM, which assumes a difference in hauling effort for moving a unit of waste vertically
versus horizontally. By determining the total ETMs necessary to move a quantity of waste
to an ex-pit location versus an in-pit location, a difference in the hauling effort could be
determined. That difference, along with a unit cost of an ETM, obtained by dividing the total
haul costs for a unit time period by the total ETMs for that time period, could then be used
to determine a total value. That total value divided by the quantity of waste (in tonnes) was
used as the estimate for the unit ton value of in-pit dumping. The formula below illustratesthe concept for the difference between hauling 100 million t of waste to a northern dump
versus an in-pit dump.
[(ETMnorth ETMin-pit) * $/ETM]/100 million t =
average unit value realised by hauling to in-pit dump versus north dump
To establish the cost of failure clean-up, records from the 1997-1998 slope failure were
reviewed. Despite extensive documentation, there is still considerable variation in what
constitutes clean-up costs. On one end of the spectrum, the costs can be merely the labor
and equipment charges associated with removing the portion of failure debris necessary to
re-establish access into a mining area or to uncover buried ore reserves.
At the other end, the clean-up costs
can include those charges along with a
range of fees associated with consulting,
additional equipment, accelerating strippingto continue mining in other parts of the
site, overtime costs, contracting and leased
equipment. Based on the previous two
assessments, a range of values was obtained
for both the unit cost of cleaning up a
tonne of failure debris and the unit value of
dumping a tonne of overburden in the pit.
Economics of in-pit dumping
The nal step was to use weighted (by
the probability of a seismic event) average
values for the expected volume of failure
debris, the expected value of the volume
of material that would have to be cleanedup, and the associated expected costs and
benets with time. Values of benets and
costs were shifted with time by using a
discount/interest rate of 7%.
Table 1 Summary of benets and costs
shifted with time
The negative values in the nal row
indicate that for the difference between
the high expected benets and low
expected costs (H/L) (best case), and the
low expected benets and high expected
costs L/H (worst case), the dump size of 30
million t is not a feasible design in this case.
This method of risk assessment has
helped Rio Tinto to understand the
interplay of a number of variables that
inuence the risk associated with placing
overburden on the north slope of the openpit. While the 30 million t dump option
proved to not be an economically feasible
option, other volumes evaluated in the
course of research do have positive values
throughout the mine life. The methodology
described here has allowed Rio Tinto
Minerals to identify those cases and
proactively manage risk in the present and
throughout the life of the project.
Ray Yost,
Rio Tinto Minerals -Boron Operations,USA
Articlereferencesareavailablefromt
heACG.
Rio Tintos Boron open pit operation was initiated in the 1950s
Open pit
YEARDUMP SIZE(TONNES)
DIFFERENCE
H/L L/H
2010 30,000,000 positive positive
2015 30,000,000 positive positive
2020 30,000,000 positive positive
2032 30,000,000 positive positive
2036 30,000,000 negative negative
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Many of the uncertainties surrounding
the development of a large open pit
mine have now been overcome with the
publication of the 496-page Guidelines
For Open Pit Slope Design.
The publication is the result of four
years of effort and support from a group
of 12 mining companies representing the
majority of the worlds production ofdiamonds and base metals.
Open pit mining is an efcient way
to mine many deposits. But there are
complications. Make the slope of the
pit too shallow and you have to move
millions of additional tonnes of valueless
overburden. But if its too steep, you risk
failure with subsequent risk to people and
property.
Up until now, the only handbook of
this type available to open pit mine slope
design practitioners, including engineering
geologists, geotechnical engineers,mining engineers, civil engineers and
mine managers has been the CANMET
manual last published in 1977.
The new Guidelines For Open Pit Slope
Design was ofcially released at the Slope
Stability conference in Santiago, Chile,
9 November. It is a direct outcome of
the Large Open Pit research project
and comprises 14 chapters that follow
the life of mine sequence from project
development to closure.
CSIRO Earth Science and ResourceEngineerings Dr John Read is one of two
editors and has also authored a number of
chapters in the book.
Dr Read has over 40 years experience
as a practitioner and consultant in the
mining industry, with special interests and
expertise in rock slope stability and open
pit mine design and investigation tasks in
Australia, Fiji, Papua New Guinea, Brazil,
Argentina, Chile, Canada, South Africa and
Zambia.
He says that each chapter is written
by an industry practitioner with specicexperience in the topic being described.
The purpose of the book is to be
a new generation guideline that links
innovative mining geomechanics research
with best practice he said. The book
outlines for todays practitioners what
works best in different situations and
why, what doesnt work and why not, and
what is the best approach to satisfy best
practice in a range of situations.
Guidelines For Open Pit Slope Design
is available from CSIRO publishing forAU$195. www.publish.csiro.au/
CSIRO helps redene largeopen pit design
Seventh Large Open Pit
Mining Conference 20102728 July 2010, Perth, Western Australia
High demand for commodities, record fuel prices and a scarcity of skilled personnel
have been replaced and surpassed by the recent global nancial crisis as the primary
issues facing the mining industry. As demand for commodities improves the incentive
to continue to drive operational and safety improvements will become paramount.
The Seventh Large Open Pit Mining Conference 2010 (LOP 2010) will provide the
opportunity to chart that progress in large open pit mines around the world
The conference will provide the forum for operations with major achievements, along
with those operators implementing changes, the chance to outline their innovations
and to share and explore experiences with others. Consistent with the aims of The
AusIMM, the Conference will allow members and the industry to keep abreast oftechnical developments and provide a forum to share views and opinions within the
large open pit sector.
For more information, please contact:
Katy Andrews, The AusIMM
Phone: +61 3 9658 6125
Fax: +61 3 9662 3662
PUBLISHING
ACG Open Pit Rock Mass
Modelling Seminar
2930 July 2010, BurswoodConvention Centre, Perth
This seminar will maximise thedissemination of geotechnical rock massmodelling and synthetic rock modelling
technologies to industry.
The trend of open pit operations
mining to steeper and deeper levels
has seen an increase in the stress
environment and greater uncertainty
about the mechanical behaviour of
slopes, elevating mine worker safety
and productivity risks. To better
identify, understand and manage thesepotential geotechnical risks (including
seismic hazard) associated with slope
stability failure, the ACG will host this
two day seminar immediately following
The AusIMMs Seventh Large Open Pit
Mining Conference 2010.
Please visit,www.acg.uwa.edu.au/events_courses
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Introduction
Mining is an important activity in
the economy of many South Americancountries. It is predominantly a formal
sector, regulated and facilitated by laws and
regulations; it is also a leading contributor
of export earnings that is integrated into
the global economy. The contribution of
the mining sector can represent up to
10% of the gross domestic product and
over 50% of the value of all exports of a
country with a strong and predominant
mining sector. Mining has a multiplier effect
- generating synergies with other economic
and social sectors in the community and
region where it was developed.
However, society does not always have agood perception of the mining industry. In
part, this may be due to the environmental
liabilities left behind by legacy mining
sites that date back to times when there
was neither awareness of the impact that
mining can have, nor a modern legal
and supervising framework. Until recently,
regulations requiring companies to prepare
abandonment and closure plans were
largely absent.
The world has changed and the
requirements for mining projects are
evolving. Compliance with internationalagreements, such as those of biological
diversity, community engagement,
climate change, and the struggle against
desertication and new environmental
standards have demanded a new way
of mining. This includes social andenvironmental impact studies and closure
plans that are developed from the time
when a mining project commences.
This article presents for comparison the
most important elements of mine closure
standards in Chile, Argentina and Peru.
Mine closure legal framework
Chile
On 7 February 2004, modications to
mining safety regulations came into force
in Chile, establishing an obligation for all
mines to prepare closure plans within
ve years. The objective is to prevent,minimize and/or control the risks and
negative effects that might result from or
continue to take place after the cessation
of the operations of a mine site, in the
life and integrity of the people working
there, and of those who, under dened and
specic circumstances, are related to the
operation and are within the inuence of
its facilities and infrastructure.
In 2009, draft law addresses the closure
scope of mine facilities and sites of the
extractive mining industry. This draft
legislation differentiates between thoseprojects that have an environmental
resolution and those that do not. The
second group are those mines that
Mine closure planning in South Americaby Hugo Rojas, Teck Resources, Chile; and Roger Higgins, Teck Resources, Canada
started operations before the Base Law
of the Environment Nr. 19300 (1997) and
Regulations of the Environmental ImpactAssessment System were enacted.
With respect to nancial guarantees,
mining companies have to provide these
in annual instalments, over a period of ve
years, or during the period of remaining
mine life (if this is shorter).
Argentina
The law on environmental protection
for mining activity and its supplementary
regulations does not contain specic
regulations for mining companies to submit
abandonment and closure plans for the
approval of authorities. This matter is opento different interpretations.
According to the Second Section of the
Complementary Title, the following must
be considered:
a) Environmental impact: modication of
the environment, whether benecial or
detrimental, direct or indirect, temporary
or permanent, reversible or irreversible,
may be potentially caused by mining
activity.
b) Environmental impact report: a
document that describes a mining
project, the environment where it isdeveloped, the environmental impact
it will cause and the environmental
protection measures proposed for
The Chilean town of Andacollo and Teck's Carmen de Andacollo mine are close neighbours. This leads to a very close relationship between the community, for bothoperations and closure planning
Mine closure
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These proceedings are a hard-bound, blackand white publication featuring 53 papers,
comprising 622 pages.
www.acg.uwa.edu.au/shop
adoption. The EIR must address
measures and actions for prevention
and mitigation of environmental
impact, and rehabilitation, restoration
or recomposition of the alteredenvironment.
c) Environmental impact declaration:
an administrative act based on the
mining environmental standards in
force, approving an EIR, passed by the
application authority, and in which are
set the specic conditions that the
holder company must comply with
during all stages of the mining project.
An aspect that is not regulated in
Argentina is community involvement in
the approval process of an EIR.
Peru
Peru applies regulations for mine closure
to every mining activity, with the purpose
of preventing, minimising and controlling its
potential risks and effects to human health,
safety, the environment, the surrounding
ecosystem and property. The regulations
were passed in 2005, and the articles
clearly specify when and what details must
be presented to the Director General
of Mining Environmental Affairs of the
Ministry of Energy and Mines.
The mine closure plan complements the
study of environmental impact and theprogramme of environmental management
corresponding to a sites operations.
The ling of the mine closure plan is an
obligation for every owner of mining
activity that is in operation, beginning
mining operations or resuming mining
operations after having been suspended
or stopped by the validity of the law,
or where there is no an approved mine
closure plan.
The approval of a mine closure plan leads
to the constitution of guarantees through
which assurance is given that the owner
of a mining activity can comply with theobligations stated in the mine closure plan.
In the event of a breach, the Ministry of
Energy and Mines can execute the closure
tasks.
An important aspect of the regulations
is the provision that allows citizen
involvement. Every stakeholder can present
their observations and make contributions.
Once the closure plan is approved, it is
to be executed in a progressive manner
during the life of the mining operation. At
operation end, the remainder of the areas,
works and facilities that, due to operationalreasons had not been closed during the
production stage must be closed. The
regulations also establish mechanisms
and periods for review, updating and
accountability.
Observations
The legal norms of closure plans inSouth America differ in their scope,
depth and citizen involvement. This leads
to different requirements for mining
operations of similar characteristics.
The review and update of closure plans
is a matter of interest for governments,
as well as for organised communities and
mining companies.
Even where there is a deciency in the
law regarding mine site closure, there are
companies that progressively design and
apply high quality closure plans.
The design of closure plans in
engineering stages prior to theconstruction of projects and their
application from the beginning of the
operations, represent an advantage
for companies and should be seen as
an opportunity to prevent, minimise
and control risks and negative effects
that might occur after the end of the
operations.
The globalisation of markets,
the requirement to comply with
international norms and standards,
the exchange and development of
technical knowledge, together with opencommunication channels worldwide,
will result in the further evolution of
mine closure regulations, both legal and
self-imposed. This will improve mining
processes and practices, environmental
stewardship and the efcient use of
resources.
The voices and actions of communities
that feel affected by mining will continue
to grow, and constructive relationships
with communities will be vital.
A good closure plan will contribute to
obtaining and maintaining the social
licence to operate.
Hugo Rojas,Teck Resources, Chile
Roger Higgins,Teck Resources, Canada
23-26 November 2010Casa Piedra Events Centre
Santiago, Chile
RESPONSIBLE CLOSURE: LIVINGUP TO COMMUNITIES AND
STAKEHOLDERS EXPECTATIONS
CONFERENCE THEMES
Designing and planning for closure
Progressive closure planning
Closure costs and nancing
Proactive stakeholder engagement
Long term water management Mine site reclamation and rehabilitation
Control and monitoring
Soil ecology
Mine cluster, redeployment,
redevelopment and decommissioning
Mine legacies and relinquishment
Legal and regulatory issues
Mining heritage and tourism
Recent closure case studies
Send your abstracts by 25 January 2010 to:
For further information, please visit:
www.mineclosure2010.com
5th International Conference onMine Closure
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Summer vacation students in winterby Peter Hills, Tasmania Mine Joint Venture, BCD Resources (Operations) NL
A phone call from Professor Marty
Hudyma in February 2009 was my
introduction to the idea of offering
summer vacation experience to students
during the winter. The concept had real
merit. We had engaged summer students
at Beaconseld before with somewhat
mixed results. This is not usually a measure
of the desire of the student to have a
go, but rather it is the coincidence of
the engagement with permanent staff
wanting to take annual leave. Inevitably,
the students are slotted in to ll the rolesof absent staff, while receiving insufcient
guidance and mentoring from remaining
staff who are left to carry the burden.
Furthermore, summer vacation students
often simply want a job to earn some
money and gain some experience. Marty,
however, was keen to see a student
undertake a project and complete real
work. The project was to be titled
Retrospective Analysis of Mining Induced
Seismicity at Beaconseld Gold Mine. It
seemed ideal. A summer vacation student
with a dened project, arriving in thewinter when minimal leave was planned
by site personnel would avoid all the usual
pitfalls of a summer placement, and so
it was agreed that a placement could be
made.
The Beaconseld Gold Mine has
experienced seismicity since 2003.
Increasing incidents of seismic events saw
the installation of a temporary seismic
array logging six uniaxial channels in
early 2004, and this was replaced by a
permanent array logging 12 channels (nine
uniaxial and one triaxial) in mid 2005. The
system was upgraded in 2007 and again in2009, and currently logs 24 channels (12
uniaxial and four triaxial).
In late 2005, the Beaconseld Gold Mine
signed on to be a minor sponsor of the
ACGs Mine Seismicity and Rockburst
Risk Management project. Sponsorship
commenced from January 2006 and has
continued since then. At the time of the
original sponsorship, the Beaconseld Gold
Mine had been experiencing signicant
mining-induced seismicity for a period
of two years. Much effort had been
expended on developing an understandingof the seismicity and procedures to deal
with it were being implemented through
the development of a Ground Control
Management Plan. The ACG software MS-RAP offered the opportunity to enhance the
management of seismicity in the day-to-day operation of the mine.
Following an accident at the mine in early 2006, all aspects of the mining operation were
redesigned under the umbrella of a Case to Manage Underground Safety (or Case for
Safety). The Case for Safety was developed in four tranches by Coffey Mining, and covered
mining of capital and operating access development (Ptzner, 2006), sill driving (Sidea, Scott
and Reeves, 2007), stoping in the generally aseismic east zone of the mine (King, Thomas
and Scott, 2007), and stoping in the seismically active west zone where the most signicant
changes were required (Scott and Reeves, 2007). A key requirement of the Case for Safety
was the establishment of protocols to manage seismicity, and MS-RAP was a key tool in that
endeavour.
Hills and Penney (2008) describe the management of seismicity at the Beaconseld
Gold Mine in some detail. Of particular utility within MS-RAP is the ability to implementOmori Analysis (Figure 1) to monitor and manage re-entry times into areas excluded after
stope blasts. Seismic analysis is coupled with intensive monitoring (Figure 2) (Penny, Hills
and Walton, 2008), including stress change using H1 cells, and the impact of that change
on the rock mass and the installed support using SMART instruments. Stope blasting is a
key trigger for stress change (Figure 3), and as a consequence it is the primary trigger for
seismic activity.
Figure 1 Omori analysis following a stope blast
Underground
Figure 2 Intensive monitoring at Beaconsfield showing the SMART cables (grey) and stress monitoring(HI cells) (yellow)
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This is especially in the west zone of
the mine where mining is conducted
remotely (Hills, Mills, Penney and Arthur,
2008), and exclusion zones of at least 50m are enforced. Other features within
the MS-RAP package are also regularly
interrogated to assist in the management
of seismicity, including the various graphical
analyses such as energy index/cumulative
apparent volume (Figure 3) and apparent
stress history, and mapping features such as
excavation vulnerability potential.
Real management decisions were being
made and inuenced by the use of MS-RAP,
but the potential of the package was not
fully realised because a signicant database
of seismic data had not been collectivelyreanalysed recently. A project was ready
made, provided somebody could be
dedicated to the task for a period of a few
months.
The student chosen to undertake
the project was Natalie Kari, a 3rd year
mining engineering student at LaurentianUniversity. While Marty provided
supervising guidance from afar, a site
based introduction to the use of MS-RAP
was provided by Johan Wesseloo, ACG.
Natalie was technically an employee of
Allstate Explorations NL during her time
at Beaconseld, and as such she technically
reported to myself.
Natalie provided the Beaconseld Gold
Mine with a substantial analysis of its
seismic data, particularly that collected
over the 18 month period to June 2009
when stoping had recommenced in earnestfollowing the 2006 accident. The database
remained live for much of her stay, allowing
Natalie to observe and understand all
Figure 4 A plot of energy index/cumulative apparent volume
the aspects of data capture through the
ISSI system, its transfer to MS-RAP, and
its analysis as an immediate tool through
Omori Analysis after stope blasts, and as a
longer term management tool in updatingEVP maps. She expended a signicant
effort in analysing data to assist in the
renement of re-entry protocols, and the
latter formed the basis of her nal report.
A synopsis of that report follows this
article. The key to understanding the basis
of a detailed data analysis such as Natalie
performed can only be gained by observing
the environment from which the data is
obtained. Consequently, Natalie went
underground to inspect the geotechnical
environment regularly, and every effort was
made to introduce her to as many facets
of mining geomechanics at Beaconseldas possible. As a result, the report she
ultimately produced has real practical
application in the ongoing management of
seismicity at the mine.
The experience of hosting a project
focused summer vacation student was a
positive one for the Beaconseld Gold
Mine. Our continued use of MS-RAP as
a tool in the management of seismicity
has been enhanced as a result. The fact
that the summer vacation student came
in the winter when vacation was not the
focus of mine staff was a signicant factorin ensuring that maximum benet could
be obtained by all parties concerned. In
particular, the benet to the students
of early career international experience
cannot be over-emphasised.
Article references are available on request.
Figure 3 A plot of raw micro-strain change data illustrating the impact of stope blasting (and nonblast-related seismicity) on the local stress field
Peter Hills,Tasmania Mine Joint Venture,
BCD Resources (Operations)NL
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Understanding mining-induced seismicityat Beaconseld Gold Mineby Natalie Kari, Laurentian University, Canada
Figure 1 Distribution of re-entry times for 73 production blasts at Beasonsfield Gold Mine in 2008 and 2009
A project was undertaken at the
Beaconseld Gold Mine to investigate the
current mining-induced seismicity at the
operation. The objectives of the project
were to identify all of the main seismic
sources currently active in the mine and to
rate the seismic sources with regards to:
Seismic source mechanism (the rock
mass failure mode causing the seismicevents).
Seismic hazard (the largest expected
seismic event that would be expected).
How mining activities (particularly stope
blasting) affects the rate of seismicity
from each of the seismic sources.
The ability for seismic monitoring to be
used as a re-entry tool for each of the
seismic sources.
The seismic analyses in this project were
all conducted using the ACGs MS-RAP
program (Mine Seismicity Risk Analysis
Program).The complex geology and geological
structures of the Beaconseld Gold Mine,
including faults, contact zones, shears,
bedding and splays, contribute to the
challenges of mining within the Tasmanian
reef. More than 8500 seismic events
were recorded at the Beaconseld Gold
Mine between March 2008 and February
2009, including nine events larger than
local magnitude +1.0. A cluster analysis
identied 56 groups of seismic events
during this period, of which 23 were
particularly active and investigated in detail.
Each group was analysed to determine theseismic source mechanism, seismic hazard
and the rock mass response to production
blasting in the mine. This analysis helped
to describe the character of each seismic
source and highlight the seismic sources
most likely to cause operational issues at
the mine. When higher hazard seismic
sources can be identied, a range of
seismic risk mitigation techniques can
be used to manage the hazard. Ten of
the seismic sources were found to have
a qualitative seismic hazard rating of
moderate-high to high. The seismic hazardrating is a good indicator of the likelihood
of larger magnitude events.
The seismic source mechanism at each
seismic source, for the one year time
period March 2008 February 2009,
was compared to the seismic source
mechanism over the last four years (June
2005 June 2009). In almost all cases, the
analysis showed that the seismic source
mechanism remained constant over time.
This is an important conclusion, as itmeans that it is the local rock mass failure
mechanism that is controlling the nature
of the seismicity, irrespective of the nearby
mining inuences. When the current
seismic response to mining is similar to the
past seismic response to mining, it gives
greater condence in using the current
seismicity to understand future seismicity.
Overall, the majority of seismic source
mechanisms at the Beaconseld Gold Mine
are related to the volumetric fracturing
associated with mining-induced stresses as
a direct response to mine blasting.An investigation of how mining activities,
particularly stope blasting, affects the rate
of seismicity from each of the main seismic
sources was conducted. The proximity
of each of the seismic sources to the
stope blasts was considered. As expected,
seismic sources in close proximity to
mine blasts have a higher rate of induced
seismicity than stopes located at further
distances. However, two particular seismic
sources did not follow this trend; often
having a disproportionately intense seismic
response to distant mine blasts. Identifying
seismic sources that do not follow
expected trends is often an indicator of
locations which have a strong geological
control. These locations require particular
vigilance with respect to monitoring and
underground inspections.
Post blast re-entry times were estimatedfor 73 production blasts, using 90% of the
total seismic energy as a re-entry criterion.
The overall distribution of re-entry times is
shown in Figure 1. Using this 90% of total
seismic energy re-entry criterion, 59 of the
production blasts had a possible re-entry
time of less than 12 hours, with 14 blasts
requiring a re-entry time of more than 12
hours. Figure 2 shows that re-entry times
are somewhat controlled by local seismic
sources and vary spatially in the mine. It
was concluded that for the Beaconseld
Gold Mine, a 24 hour re-entry period isusually conservative, although at times it
may be required. It is suggested that other
tools, such as the seismic hazard mapping
tool in MS-RAP, be used in conjunction
with the re-entry analysis when making a
nal decision on re-entry following each
blast. In addition, it is important that
this analysis procedure be continued to
monitor future changes in seismological
patterns and their potential effect on re-
entry times.
Underground
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An excavation vulnerability potential
(EVP) map was built for the Beaconseld
Gold Mine. The EVP map identies regions
of the mine that need particular attention
with regard to seismic risk management
procedures such as re-entry times,
enhanced ground support, etc. Other key
points have also come to light during the
course of this project:1. The Beaconseld Gold Mine data is
well behaved. It provides good source
parameters and locations and follows
standards and expected trends in seismic
data.
2. The seismic data gives a clear indication
of where the seismic problems are
located within the mine and where
there are no seismic problems. This is
important for future planning, and shows
that seismic monitoring is a key tool for
forecasting future problems.
3. The back-analysis shows that seismic
data identies the areas with higher
seismic hazard, or which sources are
more prone or likely to have large
events. It is apparent that some
seismic sources are more active than
others; the seismic system shows this
clearly. It is important to note that the
most seismically active sources do not
necessarily have the highest seismic
hazard.4. Daily analysis and management of seismic
data is fundamental to understanding
seismic risk.
5. At this time, the analysis did not show
any acceleration of event rate or
increased seismic hazard with depth
indicating that there are no obvious
problems with incrementally deepening
the mine.
6. It is recommended that one person at
the mine be dedicated to analysing the
seismic data and familiar with MS-RAP,
using it to its maximum potential.
It is important to note that there are
Figure 2 Location of the blasts for which the re-entry analysis was conducted
limitations to all of the analyses undertaken
in this study. Sound judgment should be
undertaken when utilising the information
provided. Focus should be placed on
minimising personnel exposure to areas ofthe mine where seismic hazard is greatest.
It is important that all available data and
tools continue to be utilised in order to
minimise the seismic risk.
Acknowledgments
This project would not have been
possible without the support, insights
and direction of several people. I would
like to express my gratitude to Marty
Hudyma, Laurentian University, and
Peter Hills, Beaconseld Gold Mine for
providing me with this opportunity and
whose supervision and direction played aninvaluable role in this project. I would like
to thank Johan Wesseloo for conducting
a site visit and help in using MAP3D. I am
also grateful to Tim Parkin, Toby Collins and
Jerome Paterson, Beaconseld Gold Mine
for providing me with guidance during the
course of my project. Notable thanks to
Roger Hill for helping me understand the
geology of the mine.
Natalies project, undertaken between
May and August 2009, was a joint effort
between Beaconseld Gold NL, Laurentian
University and the ACG. Similar student
summer undergraduate projects have been
organised each year, for the last ten years,
for sponsors in the ACGs Mine Seismicity
and Rockburst Risk Management project.
Mine Seismicity and Rockburst Risk Management Project
Natalie Kari,
Laurentian University,Canada
Since its commencement in 1999, the goal
of the ACGs MSRRM research project has
been to advance the application of seismic
monitoring in the mining industry to quantify
and mitigate the risk of mine seismicity and
rockbursting. This has seen close involvement
at research sponsor sites by undertaking
detailed site seismic analysis, testing or
experimental work and providing seismicsystem technical support and advice as
required.
Phase IV of this research project, entitled
Advancing the Strategic Use of Seismic Data
in Mines, is currently underway and aims to
develop the strategic use of seismic data and
promote an increased understanding of the
rock mass seismic response to mining. The
ACG acknowledges the generous support
and encouragement of its Phase IV research
project sponsors. Additional project sponsors
are sought.For further information please contact
project leader, Johan Wesseloo, ACG via
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The risk oftailings disposal
by Keith Seddon, ATC Williams
Introduction
In September 2010, the ACG will host
the First International Seminar on the
Reduction of Risk in the Management
of Tailings and Mine Waste in Perth. The
purpose of this article is to reect on
some of the issues that contribute to that
risk. It is written from the perspective of a
consultant.
A well known website (www.wise-
uranium.org/mdas.html) catalogues tailingsdam failures. In the (nearly) 30 years since
1980, it lists 52 incidents, spread across 20
different countries, and all continents. An
incident is broadly dened and includes
everything from contaminated seepage into
groundwater, and (relatively minor) spills
from broken pipes, all the way through
to overtopping during storm events,
catastrophic failure and collapse. The list
is by no means complete. Additionally,
inspection of the list shows an over-
representation of events from North
America, mostly related to small leaks and
spills. Are the North Americans worse atmanaging their operations than the rest of
the world? Or, is it more likely that they
are simply subject to greater scrutiny and
higher standards? These questions aside,
what can we learn from this list about the
risks of tailing storages?
Incidents occur across all mineral types.
Incidents occur across the full range of
company size and status.
Incidents occur in both developed and
under-developed countries.
The frequency of incidents does not
appear to be decreasing.
If you have a tailings dam on your site, it
is a risk.
Management of risk
Tailings storage and disposal does not
rank high on the scale of overall mine
production costs. But it does weigh
heavily in terms of the overall risk to
an operation, both initially with permits
and approvals, and in relation to ongoing
operations. There is nothing like a well
publicised tailings dam incident to damage
a companys license to operate. So,
increasingly we see that managementof mine tailings is about understanding
and management of risk. The risk based
approach is not unique to tailings storages.
It is also widely used for management of
water dams and other activities.
Two examples demonstrate the trend
with respect to dam safety. The NSW
Dams Safety Committee is currently in
the process of a comprehensive re-casting
of its requirements in order to integrate
a risk based approach. And, the ANCOLD
tailings dam guidelines (originally issued in
1999) are being updated with increased
emphasis on risk. For this approach tobe effective, a core requirement for
management is to be fully committed to
the process, through adequate support and
resources.
Fundamental hazards
There are at least four fundamental
hazards that need to be considered for all
tailings storages.
Potential energy [Gravity is a bitch]
All above ground storages place tailings
in an elevated location relative to someposition around the storage. In the event
of a breach, this potential energy may
convert to kinetic energy. This means that
the runout distances and consequences of
failure need careful consideration. These
considerations also apply to in-pit storages.
if there are underground workings belowthem.
Low strength
Strength inuences runout distances
and the assessed consequences of failure.
In addition, it also relates directly to
bearing capacity and the safe access over
the tailings for activities including raising
and capping. The strength of geotechnical
materials is tricky to dene. It varies
with time and is dependent on the rate
of loading. Almost all tailings start out as
slurries, i.e. liquid. After deposition, some
tailings progress towards the solid statefaster than others. But this does not mean
that any tailings dam can be treated like a
waste dump.
Geochemistry/acid potential
Many types of tailings contain a
proportion of sulphur, which may oxidise
to form sulphuric acid. This in turn has the
potential to mobilise trace heavy metals,
and make even small amounts of seepage
a very undesirable consequence. Little can
be done to eliminate this basic hazard; the
geochemistry of the orebody is not opento negotiation. However, in the future
possibly more attention will be given to
attempts to remove sulphides as part
of the process, and reduce the residual
hazard in the tailings. The potential for acid
production impacts both on operations
and on closure requirements for a storage.
It needs to be evaluated during the design
of all tailings dams, and may need to be
monitored routinely over the mine life.
Process chemistry
The tailings solids may prove to be
relatively benign, but it is necessaryto consider the process and how this
inuences the chemistry of the decant
water. This includes processes that use
cyanide (gold tailings), high pH (bauxite
red-mud), and low pH (laterite nickel), and
elevated levels of salinity should also be
included.
Many of the decisions relating to
process chemistry are fundamental to the
feasibility and design of the whole mine
and concentration / beneciation process,
and may be considered as constraints to
the tailings dam designer. However, whenthese conditions occur, they are likely to be
powerful drivers of the subsequent design.
The author is looking forward to the day
Tailings
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that a laterite nickel process co-locates
with a bauxite renery, and the two waste
streams are combined to neutralise each
other.
Factors contributing to risk
In risk management terminology, it is
usual to dene risk as consequence x
probability of failure. Consequence relates
to hazard, and is typically measured in
terms of loss of life, or cost of remediation.
Management of risk can address both
of these components. For instance, the
consequence of a failure will be dependent
on the location and size of a dam, and
factors such as the strength of the
contained tailings. Most of these types of
issues need to be addressed during site
selection and design. It is too late to doanything about location after a dam is built.
On the other hand, there are many issues
related to the operation and management
of a tailings dam that impact directly on
the probability of failure, whether this
is explicitly recognised during design or
not. The following discussion covers both
operation and management components,
but is slightly biased towards operational
issues.
Poor communication
Many problems stem from poorcommunication, i.e. between the designer
and site management, or management
and operators. The designer (often
a consultant) may make particular
assumptions regarding the way the dam
will be operated and raised. Typically, these
matters will be covered in a design report.
But the implementation of these lies
with the mine, and personnel rarely have
time to read design reports. A common
solution is to have an operations and
maintenance manual (OM) to cover
these aspects. A good OM manual needs
to be comprehensive, structured, wellwritten, and be easily understood by all
users. Usually the details require input
from both the designer and the mine, and
a co-operative approach to preparation is
required.
Tailings dams are not static structures,
they are continually being raised or
modied in some way, and all OMs need
to be regularly checked and upgraded to
mirror these changes.
Bad decisions [It seemed like a good idea atthe time]
There comes a time in the life of some
storages when a decision is made that
fundamentally effects safety performance,
and what can be done with the storage
in the future. This is typically something
like changing the method of raising, or the
previous water management procedures,
or the method and location of tailingsdischarge. It is not always a fully informed
or considered decision. It may be taken
under the stress of requirements to reduce
costs in the short term, or delay the
requirement for a raise, without regard to
longer term consequences. So if anyone
ever has a decision like this to make, please
put your guns back in their holsters, discuss
it with your consultant, and think through
the long term consequences, and the
implications for safety.
Water management [Soil particles and waterare natural enemies. With tailings you haveboth together]
Many operational problems associated
with tailings dams, and many of the
recorded incidents can be traced directly
to water issues. Excessive water can affect
dam safety in a number of independent
ways:
Overtopping (followed by erosion and
breach).
Increased seepage leading to piping or
internal erosion.
High seepage (phreatic) surfaces in
embankments, resulting in reducedstability.
Water balance and storm water
management in a tailings storage is usually
addressed as part of the design. Either
appropriate freeboard levels are set to
ensure that there is an adequate void space
left in the storage to retain decant and
storm water at safe levels, or a high level
emergency spillway is incorporated into
the design.
Some believe that if tailings dams contain
deleterious substances (solids and/or
process water) then provision of a spillway