Revista 2 Articulos de Diseño de Minas Sub Level Caving

8/20/2019 Revista 2 Articulos de Diseño de Minas Sub Level Caving

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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

N 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 today’s 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 Grängesberg.

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 industry’s 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


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 Kiruna Mine at three different points in time (Marklund andHustrulid, 1995)

At the Kiruna Mine today the sublevelspacing is 28.5 m. In certain sectors of

LKAB’s 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 LKAB’s 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 today’s

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 1950’s, 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


3/273Australian Centre for Geomechanics • December 2009 Newsletter

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 Grängesberg iron mine in

central Sweden in the early 1970s.

Figure 4 Results of the Grängesberg marker tests(Janelid, 1972)

Some of the relevant parameters are

summarised in Table 2.

Table 2 Design parameters at Grängesberg

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


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




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.”




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 Grängesberg, 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



Burden (m) 3

Sublevel drift width (m) 5.1

Sublevel drift height (m) 4.8Front inclination

(degrees) 75

Parameter Value




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)Grängesberg 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



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)

P e Exp

= explosion pressure for the

explosive

P e ANFO

= explosion pressure for ANFO =

1600 MPa

ρrock = 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: 70–80 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: 17–19 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 today’s 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.

rock ANFOe

Expe

hd P

P r R

ρ

65.220/ =

“The results of the marker

studies would suggest that

modications in some of

the current, very large scalesublevel caving designsshould be considered.”



Second International

Symposium on Block andSublevel Caving

20–22 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

6–8 October | Santiago - CHILE

Ponticia Universidad Católica

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 of

Mines, 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 today’s

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,

9–11 June 2008.

Figure 11 New possibilities for large-scale sublevel caving



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



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 m•3

(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.



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

A r t i c l e r e f e r e n c e s a r e a v a i l a b l e f r o m t

h e A C G .

Rio Tinto’s 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




2036 30,000,000 negative negative



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 world’s 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 it’s 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 ResourceEngineering’s 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 today’s practitioners what

works best in different situations and

why, what doesn’t 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 201027–28 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

[email protected]

PUBLISHING

ACG Open Pit Rock Mass

Modelling Seminar

29–30 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 AusIMM’s Seventh Large Open Pit

Mining Conference 2010.

Please visit, www.acg.uwa.edu.au/events_courses



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



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 site’s 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 2010 Casa 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:

[email protected]

For further information, please visit:

www.mineclosure2010.com

5th International Conference onMine Closure



Summer vacation students in winter by 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

ACG’s 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)



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



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 ACG’s 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



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.

Natalie’s 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 ACG’s “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 ACG’s 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

[email protected]



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 company’s “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



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 somethi

Revista 2 Articulos de Diseño de Minas Sub Level Caving

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