Influence of Cemented Mill Tailing as Backfill Material in Open Stopes

Influence of Cemented Mill Tailing as Backfill Material in Open Stopes

CHAPTER 1

INTRODCUTION

1.1 General Filling of the mine voids has multiple reasons such as, a simple method of tailings

disposal, or as a void filler, in a few cases it is followed as an economic method for

supporting the weak wall rocks, and lastly, for creating a working floor in a few

stoping operations. Based on the specific purpose of backfilling, the composition of

backfill material has been varied. Of the various reasons of backfilling, the role of

backfilling in providing a passive support to the weak sidewalls of a stope is gaining

importance in both cut-and-fill as well as open stoping operations. In the former, fill is

introduced periodically during the progressive extension of the stope while in the later

case, fill placement in a particular stope is delayed until production from it is

complete. In both cases, the function and duty of the fill mass can be prescribed

quantitatively. It is necessary to design the backfill to meet prescribed operational

functions and safety requirements. In order that the backfill material serves the

purposes of offering passive support to the sidewalls, cement is added as a binder to

develop strength to the backfill.

Cemented backfill became popular when it is taken as a means to support the weak

wall rock. However, the high price of Portland cement has thrown open the challenge

of economic viability. The consequence is that the researchers have tried to look for

binder alternatives which have eventually resulted in the application of high density

slurry and paste backfill materials that have improved backfill mechanical strength

response, reduced cement consumption and water disposal.

Compressive strength and permeability are the two factors affecting fill stability. The

compressive strength of fill required in mining operation varies over a wide range

depending on its application. and the variation in the strength is dependent on the

percentage of binding material.

Permeability is an extremely important fill property, which affects the stope

dewatering capability, and it should be taken into account in slurry fill design. There is

Department of Mining Engineering, IIT Kharagpur

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no designated provision for percolation rate, but the universally accepted rate in the

slurry hydraulic fill is about 2.5 to 3 x 10-3 cm/s.

The backfilling method used is often dependent on the mining system adopted and can

be classified in to two general groups; these are cyclic and delayed filling methods.

With delayed backfill, the fill must be capable of existing as a free standing wall after

being exposed during pillar recovery. From the practical point of view the stability is

more governed by the dimension of the stope in most of the cases.

Cemented fill of various solid compositions has been employed as an artificial support

with considerable popularity in the recent years. The most widely used cemented fill

consists of classified tailings, or mixtures of rock, sand and cement. Results of

practical studies show that cement content in a fill and its slurry density are essential

factors affecting fill stability and the economy of backfilling.

With accumulation of large amount of carbonaceous waste from the thermal power

plants and mill tailings from the mineral beneficiation plants of metal ferrous mines,

the environmental concern regarding the disposal of mine waste has grown in recent

years and this trend is unlikely to end. Thus the backfill technology has to address

three basic issues, namely the property of the backfill material which enhances the

stability of the ground, the utility of mine waste as a backfill material to control the

environmental pressure and lastly it has to take into account the pressures created by

the conservation of mineral resource.

The present study focus is laid on the characterization of cemented mill tailings and

attempts were made to increase the mechanical properties of the cemented mill tailings

with out adding on to the cost of production. The basic objective of increasing the

strength of the mill tailings is to study its influence on the stability of the wall rocks,

which are under severe side pressure.

1.2 Overview Backfilling of mine voids is one of the two popular techniques adopted for controlling

local, stope wall behaviour and also mine near-filed displacements. Brady and Brown

(1985) outline three mechanisms that demonstrate the support potential of mine

backfill. The three mechanisms, illustrated in fig.1.1, represent fil1 performance as


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superficial, local and global support elements in the mine structure. As a superficial

support element, fil1 provides kinematics constraint on surface blocks in distressed

rock. This helps prevent large-scale wall movements and collapse of openings. When

placed tight to the back, fill can provide back support as well. At a local scale. The

passive resistance of the fill may be mobilized by pseudo-continuous and rigid-body

displacements of the wall rock induced by adjacent mining. Lastly, fil1 may act as a

global support element and reduce stresses in the rock by carrying part of the load. In

order to provide such support the fil1 must either be very stiff or undergo large strains

in confined compression. The latter might occur in narrow-vein mining where the

hanging wall to foot wall closure strain is high.

Fig.1.1.Modes of support of backfill: (a) kinematic constraint on surface blocks in

de-stressed rock; (b) support forces mobilized locally in fractured and jointed rock; (c) global support due to compression of fill by wall closure (Brady and Brown, 1985).

The disposal of mill tailings in underground not only reduces the environmental

impact but provides the base of engineering material which can be used to improve

both the ground conditions and also acts as a working platform in a few stoping

operations. According to Barrett (1978) the purpose of the backfill is not to transmit

the rock stresses, but to reduce the relaxation of the rock mass so the rock itself will

retain a load carrying capacity and will improve load shedding to crown pillars and


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abutments. This leads to less deterioration in ground conditions in the mine, improving

operations and safety.

When used as a support medium in open stope mining. The fill must be able to remain

stable as a freestanding wall during adjacent mining. It may also be required to remain

stable during undercutting. Depending on the mining sequence and ore body geometry,

the fill may undergo various combinations of monotonic and cyclic loading prior to

and during exposure.

Adding low percentages of ordinary Portland cement (OPC) of between 3% and 6% by

weight, permits the development of cohesive strength and the ability for the backfill

mass to be self supporting when exposed in vertical faces by adjacent pillar mining.

The self supporting nature of the backfill permits higher recovery of pillar ore, which

in turn improves the utilization of the mining reserve and the economics of the mining

operation. Increased ore recovery results in longer mine lives.

In some mining methods the backfill forms a working platform for people and

equipment and therefore must be capable of supporting the traffic. Generally, cement

is not required in this application.

The placement of backfill underground directly reduces the quantity to be disposed on

surface. This has direct operating and capital cost benefits and reductions in future

rehabilitation costs.

In India, large amounts of tailings and waste rock produced annually through mining

activity. The tailings consist of a coarse fraction and a fine fraction and can be either

reactive (acid generating) or non-reactive. Most mining operations utilize the coarse

fraction for backfilling while the fines have been traditionally disposed of at the

surface in tailings ponds. The development of paste backfilling technology allows an

underground mining operation to use a portion of the fines for backfill. reduction in

tailings pond accumulation and possible elimination of impoundment for reactive

tailings in ponds and on land; reduction in costs associated with constructing and

reclaiming tailings ponds during mine start up and final close down; removal of

constraints placed upon fine grinding as an effective means of mineral value liberation,

i.e. gold, uranium and metal bearing ores.


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One practical method to modify the physical and chemical properties of the fine mill

tailings so that they can be utilized in an appropriate manner for underground

applications is by using the superplasticizer (Alsayed, 1998).

In short it can be said that backfill is one of the most important processes used

extensively in underground mining operations around the world for many purposes,

and they are:

• To provide ground support;

• To create a fresh working surface and roadways for machinery;

• To permit maximum ore recovery;

• To safe and selection extraction of ore deposits without loss of ore and

encountering dilution problems; and

• To dispose of mining waste materials.

1.2.1 Cemented Mill Tailing as Backfill Considerations There are two main types of Cemented Mill Tailing as Backfill: Hydraulic Slurry Fill,

and Paste Backfill. Each has certain advantages and disadvantages and their use is

dependent on certain aspects of the present mining method and surrounding ground

conditions. An important parameter for determining the suitability of hydraulic and

paste backfill is its strength. An adequate uniaxial compressive strength for a backfill

in a typical mine is 0.7–2 MPa (100–300 Psi), and common strength specification is 1

MPa after 28 days (Andersion, &Pigdon 2005). The paste backfill must be designed so

that it will reach its target compressive strength values at 28 days of curing age and

beyond. This can be done by choosing optimal mixtures for tailings. In backfill the

cement strength can be reduced when rock walls are sufficiently close together to help

support the backfill by shearing stresses at the wall backfill contact (Mitchel and

Smith, 1982).

An adequate modulus for a backfill in a typical mine will vary widely depending on

the surrounding ground conditions, mining method, and mining depth. Resistance of

the fill to converge should be sufficient to absorb the strain energies expected as

mining progresses.


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Hydraulic fills are slurry fills having a pulp density in the range of 60–75% solids

weight for weight. Viles and Davis (1989) state that as much as 30% of the total initial

fill volume is lost by dewatering. Hydraulic fills consist of classified coarse tailings

along with a binder. The fine tailings are usually excluded from the fill because their

removal improves flow characteristics provides better fill consolidation and

subsequent water drainage, the high water content allows the slurry to be transported

by gravity or pumping at relatively high placement rates through boreholes and

pipelines. Level preparation and clean-up can be very time-consuming with this type

of fill. The high binder dosage needed to create a hydraulic fill with good strength

properties can be expensive.

Paste fill, on the other hand, has high solids content, usually with a pulp density in the

range of 75–85% solids weight for weight. Paste backfill is cheap as comparison to

rock fill or hydraulic fill (Hassani & Archibald, 1998). This type of filling usually

contains fine material. According to Slatter (2006) as the concentration of fine particle

(below 20μm) increases, viscous stresses also increases, and paste become non-

Newtonian in nature. And it flow just like laminar. This viscous character is a dynamic

property of paste. When the paste is stationary, the attractive forces between particles

or agglomerates form a three- dimensional structure, which extends to wall of the pipe.

The shear stress required to rupture this structure and initiate flow, is called the yield

stress. Below this stress the material behaves like an elastic solid. As shear stresses and

shear rates increases, the agglomerates gradually reorientate and disintegrate, resulting

in a decrease in the viscosity of the backfill material. This process is known as shear

thinning. At very high shear stresses and shear rates, the reorientation and

disintegration process reaches equilibrium, and the viscosity becomes constant (Slater

2006).

Paste fills have gained popularity in the past few years due to several operational and

environmental benefits due to following achievement.

• The water solid ratio for the paste fill is low, producing greater strength gain

per unit volume of cement added to consolidate the fill. So strengths

approaching rock fill can be achieved, while using less cement than hydraulic

fills;


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• Facilitates a rapid mining sequence because strength is achieved earlier

compared with hydraulic fill;

• Allows the use of waste rock and slag as well as the fine fraction of tailings,

thereby reducing surface tailings impoundment requirements; so reduced the

environmental costs of the mine;

• A uniform graded backfill is capable of generating greater compressive

strength due to fewer voids;

• Paste fill has higher stiffness than hydraulic sand fill because of reduced

porosity;

• Decant water from the fill is virtually eliminated, reducing costs and

problems associated with barricade set-up/level clean-up and wear on mine

dewatering pumps;

• The present bore-hold delivery systems of slurry fills can be used;

• In saturated condition within the paste backfill the ingress of oxygen, so

limiting the potential for generation of acid mine drainage.

Due to these advantages the use of paste technology has been accepted worldwide in

the modern mining industry of today.

1.2.2 Geology of Hutti Gold Mines The Hutti Gold Mine is situated in the Archaean Hutti–Maski Greenstone Belt

(HMGB) in the south Indian Dharwar Craton (Figure1.2). The hook-shaped HMGB

consists of a heterogeneous assemblage of volcano sedimentary units, which are

metamorphosed to the greenschist–amphibolite facies transition. The western

boundary is tectonically juxtaposed to gneisses; the northern, eastern and southern

contacts are intrusive. Two generations of granitoids can be distinguished, namely the

Yellagatti and Kavital Granites. The stratigraphy of the schist belt remains unresolved

because of the lack of any definite younging direction criteria and radiometric dating.

A granitoid pebble within a conglomerate forming the base of the HMGB has yielded

a SHRIMP weighted mean 207Pb/206Pb zircon age of 2576F12 Ma, which is

interpreted as the maximum age of the belt.


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Fig.1.2. Schematic geological map of the Hutti– Maski Greenstone Belt (Jochen Kolb 2004). Inset shows location of the study area in South India.

1.3 Need of the present Investigation Several mines in India are using hydraulic backfill, cemented rock filling etc., but until

now there in case where paste backfilling is being used. Paste backfilling is far better,

cheap than hydraulic backfill. The published research work in India has been on the

preparation and transport of hydraulic backfil1 rather than improving its cured

geotechnical properties. However in the last two decades a considerable extent of work

has been reported to have been carried out in many other countries on the aspects of

improving the strength and flowability of concrete by adding superplasticizer (Alsayed

98, Yoshioka 2002, Nkinamubanzi 2004, Khatib 1999, and Radocea 1992). However

the effect of superplasticizer on strength and flowability in cemented paste backfill

was not attempted. In the absence of any detailed laboratory work on paste fill strength

and deformation behaviour, in the present project work, the initial design of cemented

backfill technology for The Hutti Gold Mines Ltd. has laid its focus on improving the


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strength of the backfill material, with different proportions of cement as well as the

addition of superplasticizer.

1.4. Objectives In order to develop an understanding of the strength and deformation behavior of

Cemented mill tailing as backfill and to optimize its use at the Hutti Gold Mine. The

specific objectives of this thesis are to:-

• Review the pertinent physical and mechanical properties of paste backfill

along with relevant aspects of soi1 mechanics and published results of

laboratory testing Paste backfill;

• Laboratory testing is identify a cost – effective backfill mixture which will

fulfill the desired strength and deformation behavior of Hutti gold mine

Cemented mill tailing backfill as a function of binder content and cure time

in uniaxial, direct shear test. So that the mix characteristics will be adjusted

in such a way that when underground opening is filled with this mixture,

the filled structure will safely withstand strata loading, and will limit

underground and surface movements; and

• To develop an understanding of the performance of cemented paste backfill

when exposed to superplasticizer.


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CHAPTER 2

LITERATURE REVIEW

The literature review is focused on the published results of laboratory testing of

backfill material. It is structured around specific geotechnical properties and includes

discussion of relevant aspects of established soil behavior. A vast amount of literature

exists on laboratory testing of conventional hydraulic fill. Since paste fil1 is a

relatively new technology, most of the existing literature focuses on preparation and

transport rather than geotechnical characteristics. The following aspects of paste

backfill behaviour will be critically reviewed:

I. Specific gravity;

II. Particle size distributions;

III. Porosity & void ratio;

IV. Permeability;

V. Unconfined compressive strength;

VI. Shear strength.

VII. Superplasticizer

2.1 Specific Gravity The specific gravity of a fill is the ratio the weight in air of a given volume of fill

particles to the weight in air of an equal volume of distilled water. The specific gravity

of a fill is often used in relating a weight of fill to its volume. By the help of specific

gravity we can find the unit weight of fill. And unit weight is very useful for finding in

nearly all pressure, settlement, and stability problem in backfilling.

2.2 Particle Size Distributions The size distribution has the largest effect on fill porosity and delivery. Tailings have

been the most widely used materials in backfill, and as such their properties have

become of main interest. It is the fine fraction of the tailings that makes a fill into a

paste. A paste fill contains at least 15 % by weight of particles less than 20µm (Slatter

2006; Amaratunga 1997). It provides the water retention properties needed to prevent


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water bleeding, prevent One of the most important characteristics of any fill material is

a well-graded particle size segregation of larger particles in the fill, and act as a

lubricant during pumping (Amaratunga 1997).

A backfill should have a suitable particles grading for any of its coarse aggregates.

Viles and Davis (1989) state that suitable backfill materials have similar features to

those specified for concrete aggregates. If coarse aggregates (≤ 25μm) are incorporated

into the mix, it must have a well graded distribution to help reduce the void ratio in the

fill. Particles size distribution curve can be described by the coefficient of curvature

(Cc) which should be between 1 and 3 for a well graded distribution (Das 2002).

2.3 Porosity & Void Ratio Porosity is defined as the amount of air space in a soil that is available for fluid flow

(Meegoda and Gunasekera 1992). The porosity, η, defined as the ratio of void volume

to total volume in a fill mass.

While, the void ratio, e, is defined as the ratio of void volume to solid volume in a soil

mass,

So, ( )(%), 100( )volumeof airporosity nvolumeof soil

= × 2.1

(1 )en

e=

+ 2.2

Porosity decreases markedly with pouring pulp density within the practical hydraulic

range of 65- 75% pulp density. At low pulp densities a porous cement gel is formed

(Mitchel and Wong, 1982) which in turn, produces higher backfill porosity. The

porosity of a fine tailing decreases with addition of cement

Many author present the results of an investigation on the influence of porosity (n) on

the uniaxial strength σn of various engineering materials. Fig. 2.1 shows the results of

test work by Li. Li; and Aubertin. M (2003) has determined the effect of each of these

factors On UCS. Higher UCS from lower porosities due to greater particles

interlocking and the fact more cement is available per unit volume of backfill.

Porosity is generally easy to measure with an acceptable level of accuracy, so it is a

practical property to use to characterize the influence of material internal structure. It


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should also be kept in mind, however, that porosity, taken as an average scalar

parameter, provides only part of the information.

Fig.2.1. The effect of porosity on UCS of cemented sand (Li. Li. et al. 2003).

2.4. Permeability Permeability is the measure of the ability for pore fluids to pass through the fill

material. Permeability is the key to successful drainage. Upon placement it is desirable

to have a fill drain as quickly as possible so that the fill material can settle and reach its

operating density. This process of drainage is dependent upon the permeability of the

fill. Thus preparation of the backfill is often required to remove the fine particles.

Paradoxically the best result strengths are obtained at high binder levels where the

binder is as fine as possible, therefore reducing permeability (Viles and Davis 1989).

Viles and Davis (1989) has given one solution to these problems is to eliminate the

need drainage by locking up the water within the backfill.

The importance of permeability can be seen in the fact that hydraulic mine backfill

must be dewatered from a moisture content of 30-40% (by weight) to around 20%

before being able to perform its ground support role (Viles and Davis 1989).The

processes by which dewatering occurs are decanting, exudation, and percolation, all of

which are dependent upon the permeability of the fill.

It is common practice to specify fill permeability in terms of percolation rate.

From Darcy’s Law the coefficient of permeability can be defined as (Aysen 2002)


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hALQKΔ××

= 2.3

Where K= the coefficient of permeability (m/sec)

Q= the volume rate of flow through the fill (m3/sec)

L= length of fill (m)

A= the cross-sectional area of the fill (m2)

Δ h = the head differential across the fill (m)

The percolation rate is the same as the coefficient of permeability except that the

hydraulic gradient hLΔ is equal to unity. Therefore, the percolation rate of any fills

material can be written as:

Kp=Q/A 2.4

Where Kp = the percolation rate of fill (m/sec)

Since the percolation rate is dependent on the fill material, the grain size distribution

and the porosity affect the drainage behavior of the fill. As for the fluid flowing

through the fill, in this case water, the viscosity and degree of saturation of the fill also

influence its dewatering capabilities.

The permeability coefficient varies considerably with grain size, Plasticity, and void

ratio demonstrate that Hazen's formula is best for predicting the average permeability

of sand tailings(Rankine 2002)

K=C d102 2.5

K= permeability constant, cm/sec

Where d10 = grain size in millimeters for which 10% of the particles pass by weight.

C=constant (≈1)

For a given size distribution, permeability will decrease with void ratio. A 20%

decrease in void ratio can cause a 50% decrease in permeability. The addition of

cement to tailings also decreases the permeability of tailings with finer materials

experiencing a greater percentage decrease (Mitchell and Smith 1982). The effect of

cementing reactions is to reduce the porosity of the fill and block drainage paths.

Permeability is an important design parameter for slurry fill. A good permeability is

required to ensure that excess transport water drains from the fil1 as quickly as

possible. The general industry accepted standard for minimum permeability of slurry


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fill is about 2.5 to 3 x 10-3 cm/s determined using either a constant head or falling head

permeability test.

2.5. Unconfined Compressive Strength The unconfined compressive strength (UCS) is regarded as the most useful measure

for comparing the relative strengths of paste backfills. There are several factors which

affect UCS including cement content, water-solids ratio, size distribution, and cement

type. Although the cement content contributes directly to the strength of the fill, the

final amount of cement is controlled more by economics than by design strength

parameters.

Laboratory work has also indicated that the grain-size distribution of a fill material is

important in optimizing its strength. Modification of a fill’s grain-size distribution is a

useful tool for improving a fill’s strength characteristics. The “desirable” size

distribution can be described as well-graded, containing course to fine particles

Thomas (1973) indicated that as the fines content of a fill material was increased, the

strength increased. Three cemented fills, prepared from deslimed tailings and

distinguished according to their grain-size distributions, were tested for strength.

Although comparable, the strengths of the medium and coarsely graded fills are less

than that of the well-graded fill that contains more fine particles.

2.5.1. Effect of Curing Conditions The normal curing procedure for paste backfill, outlined by Archibald (1997), is to let

the sample drain for 48 hours then set it aside in a sealed polythene bag. Many paste

fills with a measurable slump are prepared in an oversaturated condition. If this is the

case, the paste fil1 may produce some decant water or bleed water from the base of a

sample tube which is not completely sealed. If the amount of water is small (e.g. 1-2%

of total sample weight), the effect of letting the water bleed from the sample prior to

curing should not be significant. The conditions under which the sample is cured have

a much greater effect on sample strength. If the sample is allowed to dry out at all, the

UCS will increase since lower moisture content will result in better cement hydration


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(Mitchell and Wong, 1982). This is not considered realistic for paste fil1 because it is

not free-draining and remains close to saturation for a much longer period of time.

The curing time, temperature, and humidity have also been shown to have an effect on

cemented fills. As with normal concrete, the strength of a cemented fill increases with

curing time. Perry (1990) tested specimens of differing water/cement ratios and found

that ultimate fill strength is achieved after long periods of up to 90 days. Elevated

curing temperatures also accelerate the gain in strength of cemented backfills provided

that full saturation is maintained and the temperature does not exceed a critical level.

Kejin Wang et al. (2004) observed that strength can be increased by increasing curing

temperature and found that for common backfills, a curing temperature of 50°C is a

safe limit. As opposed to time and temperature, the strength of cemented fill decreases

with humidity. Compressive strength increases more rapidly when dry cured rather

than wet cured (Mitchell and Wong, 82). Curing temperatures less than 15° C retard

reactions while higher temperatures accelerate reactions, excessively high curing

temperatures degrade humic materials in highly organic soils, detrimentally affecting

strength gain (Jesse Jacobson 2002)

Although these factors affect the strength of cemented fills, they may be impossible to

control due to the underground environment. Whatever curing temperature is used, the

specimens should be properly spaced and fans or pumps should be used to ensure that

all specimens cure at the same temperature (Jesse Jacobson 2002). It is important to

note that, excluding curing time, few mining environments will be able to provide the

ideal hot and dry environment that can help improve a cemented fill’s strength.

G. M.M. Ley, C.M. Steed, D. Bronkhorst and Robert Gustas (1998) have found in the

case of the garson mines that the laboratory material values (average UCS .65 MPa

with an average Young’s Modulus of 215 MP) for the paste fill varied significantly

from those obtained from insitu samples, with the insitu values being stronger and

(1.5MPa) stiffer (920MPa). Temperature monitoring carried out during fill placement

indicated high hydration temperatures which may account for the increased strength.

Although humidity control is not standardized, Jesse Jacobson (2002) recommends

several methods for controlling the humidity in the curing environment: curing

samples in sealed, airtight tubes; curing underwater; or placing samples inside an

insulating jacket found that providing the samples access to water while applying a


15


confining pressure during curing, which may imitate field conditions more accurately,

reduces strength.

Although free water in sample will readily evaporated at a relative humidity less than

45%, the cement gel water appears to resist evaporation and cured water content, under

low relative humidity, increases with increased cement content (Mitechell and Wong,

1982). Humid cured specimen, on the other hand, showed a general deceases in water

content with increases cement content, a result due to Hydration. Uniaxial strength, for

given cement content, was found to increase as the water content decreased.

2.5.2 Pulp Density Pulp density is a measure of the percentage of the solids by total weight in the slurry or

paste backfill. Pulp density is an important factor in fill strength development and

pumping characteristics due to water content paste fills have a pulp density which

range from 75 % to 88% while hydraulic fills are placed at pulp density of 55% to 75

%( Amaratunga, and Yaschyshyn, 1997). For strength purpose a high pulp density is

ideal, however, paste fill pumping and placement consideration limit the density.

Fig. 2.2 shows the results of test work by Benzaazoua, (2003) have a significant

influence on the mechanical performance of the paste fill in different ways depending

on their concentration in the pore water. Fig.2.2 clearly shows that decrease in strength

(UCS) occurred with an increase in the water content (or slump value). This effect is

clear at short-term (7 days of cure) as well as medium term curing times (28 and 56

days of cure), however it is less clear at the long-term curing time (120 days of cure)

possibly due to a certain internal weathering of the samples as seen for samples made

of tailings


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Fig.2.2. Effect of the water content (w%) on UCS (M. Benzaazoua 2003)

2.5.3 pH When a significant quantity of lime is added to a mill tailing, the pH of the mill tailing-

lime mixture is elevated to approximately 12.4, the pH of saturated limewater at 25° C.

This is a substantial pH increase for mill tailing. Mehta and Monteiro (1993) show that

the solubilities of silica and alumina are greatly decreased at reduce the pH levels,

which can lead to a decrease in pozzolanic reactions as well as decreased cation

exchange capacity. Mehta (1993) suggested that the high pH causes silica from the

clay minerals to dissolve and, in combination with Ca++, to form calcium silicate. This

reaction will continue as long as Ca (OH)2 exists in the mill tailing and there is

available silica. A very small amount of Ca (OH)2 was required to raise the pH to the

target value.

2.5.4. Mineralogical Effects The mineralogy of the mine tailings and cement plays an important role in the strength

development of cemented paste backfill. The mineralogical composition is determined

by X-ray diffraction analysis (XRD), which provides the crystalline mineral

assemblage of a sample.

It is common knowledge that sulphide mine tailings generate sulphuric acid in the

presence of water and oxygen, which may lead to possible chemical weathering and

other consequences. The presence of sulphide minerals within cemented composites as


17


well as the soluble sulphates has a deleterious effect on the strength of paste backfill

due to sulphate attack (Kesimal, & Ercikdi, 2004).

The effect of sulphate on the strength of cemented paste backfill depends on the

strength of sulphate concentration, the curing time, and the cemented composition and

content (Benzaazoua et al. 2003).

According to Archibald (1999) the deleterious nature of sulphide on cement

performance increases with increasing O2 concentration, specific surface area and

temperature. pH is also an important factor in sulfide oxidation. At low pH, bacterial

action is an important factor in oxidation. But this role is eliminated at pH levels above

approximately 4. A typical sulfide oxidation reaction for pyrite, described in the

following equation illustrates the reason for its harmful effect on cement hydration.

FeS2 +15/4 O2 +7/2O2→Fe(OH)3(s) +2SO42- +H+ 2.6

the strength loss experienced by cement, curing the presence of this reaction, is due to

the production of hydrogen ions causing an acid attack that dissolves the calcium

hydroxide found in hydrating cement and the precipitation of gypsum, ettringite and

monoaliminate sulphate by the reaction of aluminates in the cement the mineral

species have low molar densities compared to the cement components they replace and

thus, cause expansion in cements.

Mitchell and Wong (1982) note that large variations in strength should be expected

between cemented tailings sands of different mineral or chemical composition. They

performed tests on tailings with 2.5% sulphur content and found that it produced

strengths that were only 60% of the strength of non-sulphide tailings prepared with the

same cement content due to an internal sulphate attack, which results from the

chemical interactions of the sulphate ions with the Portland cement hydration products.

They indicate that the possibility of long-term strength loss in some cemented tailings

backfills should be considered in backfill design and mine operations.

Kesimal (2004) also find that effect of sulfide on UCS. It was pointed out that for high

sulphide tailings, neither the slag-based binders nor the fly-ash based it was pointed

out that for high sulphide tailings; neither the slag-based binders nor the fly-ash based

binders were effective, whereas the sulphate-resistant-based binder gave good long-

term strength. To control the deteriorating effects of sulphate attack, the amount of


18


calcium hydroxide and calcium aluminates hydrate must be kept to a

minimum(Hassani et al. 2001).

2.6. Shear Strength

2.6.1. The Mohr –Coulomb Failure Criterion The shear strength of backfill has traditionally been represented by the Mohr-Coulomb

failure criterion which States that shear strength increases with increasing normal

stress on a plane:

τ=c+ σ tanΦ 2.7

Where τ = shear stress on the failure plane.

c = cohesion intercept.

σ = normal stress on the failure plane, and

Φ = angle of shearing resistance.

Since shear stress cm only be resisted by the skeleton of solid particles according to

Terzaghi’s effective stress law. Shear strength is expressed as a function of effective

normal stress:

τ=c'+ (σ –u) tanΦ' 2.8

τ=c'+ σ' tanΦ' 2.9

Where u= pore pressure, and

c' and Φ' = shear strength parameters in terms of effective stress (drained shear

strength parameters).

2.6.2. Effective Shear Strength The determination of the internal friction angle (φ) and the effective cohesion (c) is

commonly accomplished by the direct shear test or the triaxial test. The direct shear

test is preferred because of its simplicity and lower cost. The advantages and

disadvantages of direct shear tests are given below.

Advantages of direct shear testing are as follows:

• The test is relatively inexpensive and quick to perform.

It requires less sophisticated equipment than other methods and it is easier to reduce

the data and interpret the results


19


• It has been found that soil parameters φ and c obtained by direct shear

testing are nearly as reliable as triaxial values. Typical values obtained

with the direct shear test are 1 to 2 degrees larger than values obtained

with the triaxial test (Bowles, 1979).

• It is good for measuring residual strength values.

Disadvantages or limitations of the direct shear test (Holtz and Kovacs, 2003) are as

follows:

• Shearing stress is not uniformly distributed across the sample. Initial

failure occurs at the corners and ends of the box, and propagates

towards the center.

• The test forces failure to occur along a fixed zone or plane.

• There is an uncontrolled rotation of principal planes and stresses that

occurs between the start of the test and failure.

• Pore water pressures for fine-grained soils are neither controlled nor

monitored.

2.7 Superplasticizer

2.7.1 Backfill Strength Enhancement by Superplasticizer

The properties of backfill material are governed by its strength and flow behavior,

which can achieved by the dispersion of backfill material. It is widely known that

better strength and fluidity is achieved by the addition of superplasticizer (Chandra,

2002.).

A superplasticizer is one of type of admixtures called water reducers (Stuart.1980) that

are used to reduction in water requirement of concrete. Water reduction results in

undesirable effect on setting, bleeding, segregation and hardening characteristics.

Superplasticizer is chemically different from normal water-reducers, and is capable of

reducing water contents by about 30%. The presence of superplasticizers (SP) in a

concrete mixture is quite advantageous, in that they assist in the effective dispersion of

cement particles due to the electrostatic repulsion (Termkhajornkit 2004) and yield

stress value decreased but the plasticity does not decrease significantly, paste will


20


obtain good flow ability without ingredient segregation. So it improves the workability

of concrete (V.S. Ramachandran, 1979 and P. Termkhajornki, 2004). The water to

cement ratio is reduced when SP is added to cement paste, which leads to reduced

permeability, increased strength and producing durable concrete

Due to these excellent properties superplasticizer can use in cemented paste backfill.

So basic advantages of superplasticizers in Cemented Paste Backfill (CPB) are -

(1) High workability of CPB, resulting in easy placement without reduction in cement

content and strength; (2) high strength CPB with normal workability but lower water

content (fig. 2.3); and (3) a binder mix with less cement but normal strength and

workability (4) CPB of good workability has advantages in that it permits easy and

quick placement. (5) It increases the flow characteristics and becomes self leveling.

.

Fig. 2.3. Effect of Super plasticizer on Strength (Stuart, 1980)

Sokalan HP 80 is an advanced superplasticizer that consists of a polycarboxylate. It has

the following advantages.

• It creates strongly plasticizing effect on cement for high strength CPB;

• Longer working time; and

• Low consumption compared to conventional plasticizers.

Khatib (1999) states the inclusion of superplasticizer in cement paste leads to a

reduction in the total pore volume and to a refinement of the pore structures. The


21


dominant pore size is unaffected and the threshold diameter is reduced in the presence

of SP. The initial curing regime has a substantial influence on porosity and pore

structure. Less pore volume and finer pore structure are obtained when cement paste is

subjected to initial moist curing as compared with initially dry curing (Khatib, Mangat

1999). Flow velocity of cement paste increases with the increased concentration

(Yoshioka 2002).

2.7.2 Rheological Test The transportation of cemented mill tailings in the form of paste through pipelines is

one of the main stages of paste backfill operations. One of the data-sets used for

pipeline design purposes are those correlating the yield stress of fluid material changes

with changes in friction loss and the diameter of pipes (Li and Moerman, 2002).

Furthermore, the yield stress-water content curves are usually used in the design of

pumping energy requirements for the transportation of paste backfills through

pipelines (Van Dijk, 2001); (Sofra and Boger, 2000). Therefore, the yield stress

measurement is of great importance to understanding the pipeline transportation

principles of paste backfill.

The yield stress of viscoplastic fluid material is defined as a critical shear stress at

where it begins to flow as viscous material with finite viscosity. Within the context of

rheology, paste backfill is categorized as a viscoplastic fluid; a non-Newtonian,

homogenous mixture that may show time-independent or time-dependent behaviour

(Cooke, 2001); (Brackebusch, 2002); (Li and Moerman, 2002). In either case, the yield

stress depends on the characterization of the mixture (Nguyen and Boger, 1992).

2.7.3. Apparatus The yield stress of the specimens can find by using the vane method in controlled

shear rate (CSR) and controlled shear stress (CSS) modes, and slump tests by cone or

cylinder (Clayton 2003).

In the slump test methods, standard conical and cylindrical slump moulds can use.

Clayton (2003), Saak (2004) has find that the cylindrical model is an excellent


22


technique for slump test because its value is fully agreement with the experimental

yield stress data obtained using the vane method.

According to Clayton (2003) the cylindrical slump test has many advantages over the

cone slump test:

• Cylindrical test value is fully agreement with the experimental yield stress data

obtained using the vane method.

• The cylinder model is mathematically simpler which is important for operators

who may not have a strong background in mathematics.

• Due to the more complex cone geometry, the cone is more difficult to fill,

during experiment; full extraction of material is also very difficult. Due to this

cause leading to the likely presence of air bubbles which can adversely affect

the results.

• Cylinder slump measurements can be completed with a section of pipe or even

a beer can, whereas cone measurement must be completed with a cone

manufactured to certain specifications.

2.7.4. Theory The cylinder model is generalized for any-sized cylinder. The cylinder theory

developed by Pashias (1996) is also presented to enable easy comparison with the

generalized cone theory. Schematic diagrams for the cylinder are presented in fig. 7.

The schematics display the important variables and the stress distributions involved in

slumping.

Fig.2.4. Schematic diagram of the cylinder slump test, showing initial and final stress distributions (developed from Pashias et al., 1996). The dimensionless yield stress of backfills can be calculated from slump height, s,

using the following equations (Pashias et al.1996):


23


'0

1 h'2yτ = 2.10

)]2ln(1[21s yy τ′−τ′−=′ , 2.11

Where τ'y is the dimensionless yield stress

' yy gh

ττ

ρ= 2.12

s' is the dimensionless slump

' ssH

= 2.13

h0' is dimensionless height of undeformed region,

' 00

hhH

= 2.14

h1' is dimensionless height of deformed region.

' 11

hhH

= 2.15

and H is the height of cylinder.

The actual yield stress can then be estimated by multiplying τ'y with ρgH.


24


CHAPTER 3

LABORATORY INVESTIGATION

3.1. Experimental Method

3.1.1 Specific Gravity The specific gravity of backfill material was carried out in 25 ml bottles (also known

as pycnometer bottle). The density bottle with fill material, half full with water, were

placed in warm water bath for 30-45 minutes to remove entrapped air. This was latter

placed in a desecrator where a Crompton Parkinson vacuum pump, of 0.25 KW, 2.3

amp, rpm 1425 was used for about an hour to completely remove any remaining air

within the sample.

3.1.2. Particle Size Analysis

Particle Size Analysis is performed to evaluate the particle size by sieving procedure

most appropriate to the samples from Hutti gold mines. Generally, the distribution of

particle sizes larger than 75 μm is determined by sieving.

Following sieves was used for particle size analysis is numbers: 4, 16, 25, 40, 60, 70,

100, 140, 200, and 270

3.1.3 Porosity 50 gram mill tailing with different cement % taken in different measuring cylinder

and volume was measured. Then transfer the content of different cylinders in other

cylinders which are contains 50 ml water. Now leaved the composition in cylinder for

settle 1 day and then record the combined volume of the MT composition and water.

And measured the porosity by formula 2.1

( )(%), 100( )volumeof airporosity nvolumeof soil

= ×

3.1.4. Permeability Using the constant and falling head permeability tests the coefficient of permeability

was determined. Initially water flows through the sample until flow (Q) and the


25


hydraulic head loss (∆h) has reached a steady state. The flow rate and head loss are

then measured and the coefficient of permeability calculated by equation 2.3:-

hALQKΔ××

=

Where, ‘Q’ is the flow rate (q/t), ∆h is the constant head loss and, ‘A’ and ‘L’ are the

cross sectional area and length of the sample respectively.

3.1.5 pH

Air-dried soil samples were adjusted to 100% water content by adding distilled water,

and pH values were measured using a calibrated pH probe.

3.2. Uniaxial Compression

3.2.1 Sample Preparation The purpose of the uniaxial compression tests was to obtain unconfined compressive

strengths (UCS) and moduli as a function of binder content and cure time. Al1 test

samples were cast in the laboratory. The different cement contents were sampled for

each type of test: 3%, 6%, 10% and 20% by dry weight (Cement: mill tailing). In all,

60 samples were cured on laboratory for 3, 7, 14, 21 and 28 days in summer season.

Other samples were cured on laboratory for 28 days for different pulp density in winter

season. Again 24 samples were cured on laboratory for 28 and 90 days for different

composition of superplasticizer in winter season.

The samples of 54 × 110 mm diameter by length were cast in the Department in

wooden molds (Fig.3.1). After allowing them to set for 48 hours, al1 of the sample

were removed from the wooden molds and were waxed at both ends to prevent

moisture loss due to evaporation and possible oxidation of the samples. The samples

were cure for 3, 7, 14, 21 and 28 days

Fig.3.1. Sample preparation in the wood mold

3.2.2 Uniaxial Compression Test


26


Both ends of the samples initially were done parallel by polish. Before the testing

samples length, diameter and weight were measured. The sample was placed in the

testing frame its stroke control rate was 0.315mm/min and brought into contact with

the load cell by adjusting the hydraulic ram (fig.3.2). When the sample was failed load

and deformation was noted. UCS was calculated with the Secant value of Young’s

modulus at 50% peak stress (Unal 2002).

Fig.3.2 Uniaxial Compression test machine

3.3. Direct Shear Test under Consolidated Drained Conditions

The direct shear box of 50 × 50 × 25 mm was used for conducting shear strength

tests. All the direct shear tests were conducted in the laboratory using manual operated

equipment. Prior to testing, all samples were air-dried. All of these samples were

tested under five different values of normal stress from 19.6 to 76.6 KPa and at a

shearing rate of 0.25mm/min.


27


CHAPTER 4

LABORATORY TEST RESULTS

Several tests were performed on cemented Hutti Gold mill tailings to evaluate the

basic material properties that influence the strength and deformation behaviour of

paste fill.

4.1 Specific Gravity Specific gravity has found 2.67

4.2. Mineralogy and Chemical Composition A qualitative assessment of tailings mineralogy using X-Ray diffraction (XRD)

indicated the mine tailings consists mainly of quartz (SiO2), followed by Albite

(NaAlSi3O8), Calcium Peroxide (CaO2), Cordierite(Mg2Al4Si5O18), Potassium sulfate

oxide(K2S2O5), sulfur (S7)and Sodium Manganese Silicate (Na2Mn6Si7O21) as shown

in the Fig. 4.1. The relative proportions of the minerals are based on peaks from an X-

ray diffraction analysis of the Hutti gold mine tailing. Chemical composition had

determined by scanning electron microscope method given in Table 4.1.

The mineralogy analysis shows that sulphide minerals are not present. And its pH

value is 7.89. So strength will be not more affected due to bacterial action in oxidation.


28


P o s i tio n [°2 The ta ]

2 0 3 0 4 0 5 0 6 0 7 0 8 0 9 0 1 0 0 1 1 0

C o unts

0

1 0 0

4 0 0

9 0 0

1 6 0 0

Mg2

Al4

Si5

O18

Ca

Mn

Si;

Ca

Mn2

As2

K2

S2

O5;

Mg2

Al4

Si5

O18

; S7

Si O

2; M

g2 A

l4 S

i5 O

18N

a A

l Si3

O8;

Mg2

Al4

Si5

O18

; S7

K2

S2

O5;

Na

Al S

i3 O

8; C

a M

n S

i; C

a M

n2 A

s2; S

7N

a A

l Si3

O8;

Ca

Mn

Si;

Ca

Mn2

As2

; S7

Si O

2; N

a A

l Si3

O8;

Mg2

Al4

Si5

O18

; S7

Na

Al S

i3 O

8; C

a M

n2 A

s2; S

7K

2 S

2 O

5; N

a A

l Si3

O8;

Mg2

Al4

Si5

O18

; S7

K2

S2

O5;

Mg2

Al4

Si5

O18

; S7

K2

S2

O5;

Na

Al S

i3 O

8; C

a M

n S

i; S

7

K2

S2

O5;

Ca

Mn

Si;

S7

K2

S2

O5;

Na

Al S

i3 O

8; C

a M

n2 A

s2; S

7K

2 S

2 O

5; S

i O2;

Na

Al S

i3 O

8; M

g2 A

l4 S

i5 O

18; S

7

K2

S2

O5;

Si O

2; N

a A

l Si3

O8;

Ca

Mn

Si;

Mg2

Al4

Si5

O18

; S7

K2

S2

O5;

Si O

2; M

g2 A

l4 S

i5 O

18; S

7

K2

S2

O5;

Si O

2; N

a A

l Si3

O8;

Mg2

Al4

Si5

O18

; Ca

Mn2

As2

; S7

K2

S2

O5;

Na

Al S

i3 O

8; M

g2 A

l4 S

i5 O

18; C

a M

n2 A

s2; S

7

K2

S2

O5;

Na

Al S

i3 O

8; S

7K

2 S

2 O

5; S

i O2;

Na

Al S

i3 O

8; C

a M

n S

i; C

a M

n2 A

s2; S

7

K2

S2

O5;

Si O

2; C

a M

n S

i

K2

S2

O5;

Si O

2; C

a M

n S

i

Si O

2; C

a M

n S

i; C

a M

n2 A

s2

Si O

2; C

a M

n S

i

K2

S2

O5;

Si O

2; C

a M

n S

i

K2

S2

O5;

Si O

2; C

a M

n2 A

s2S

i O2;

Ca

Mn

Si;

Ca

Mn2

As2

Si O

2; C

a M

n S

i; C

a M

n2 A

s2

Si O

2

Si O

2

Si O

2

TA IL IN G b .rd

Fig.4.1 X-ray diffractogram of hutti gold mill tailing

Table 4.1 Chemical Composition of Hutti gold mill tailing (determined by Scanning Electron Microscope method)

Chemical component % by weight Element % by weight Na2O 5.88 Na 6.72 MgO 6.23 Mg 6.20 Al2O3 9.83 Al 8.97 SiO2 42.31 Si 35.55 SO3 4.89 S 3.79 K2O 1.1 K 1.67 CaO 9.08 Ca 11.76 TiO2 1.46 Ti 1.58 MnO2 .34 Mn .47 Fe2O3 17.43 Fe 21.29 NiO .1 Ni .13 ZnO 1.35 Zn 1.89

4.3. Particle Size Distribution The particle size distribution of the tailings was measured using a sieve analysis (table

4.2). Fig. 4.2 shows the size distribution obtained by sieve analysis method compared

with a size distribution obtained by sieve analysis:

4.3.1 Sieve Analysis Mass of dry mill tailing taken was 500 gram

Mass loss during sieve analysis is given by the following equation


29


1( )% W WMass lossW

100−= × 4.1

(500 498.6)% 100 0.28%500

Mass loss −= × =

Table.4.2. Sieve Analysis result

Sieve

no.

Opening

size

(mm)

Mass of

soil retained

on each sieve

Wn

% of mass

soil retained

on each sieve

Rn

Cumulative percent

retained ∑Rn

Percent

finer,

100 -∑Rn

4 4.75 0 0 0 100

16 1.18 1.6 0.32 0.32 99.68

25 0.71 1 0.2 0.52 99.48

40 0.425 2.1 0.42 0.94 99.06

60 0.25 8.7 1.74 2.68 97.32

70 0.212 21 4.2 6.88 93.12

100 0.15 80 16 22.88 77.12

140 0.106 266.3 53.26 76.14 23.86

200 0.075 58.1 11.62 87.76 12.24

270 0.053 59.1 11.82 99.58 0.42

Pan

0.7

0.14 99.72 0.28

∑=498.6


30


0.1

1

10

100

0.01 0.1 1 10

Grain Size D(mm)

% P

assi

ng

Fig.4.2 Plot of Percent Finer vs. Grain Size

The following properties were obtained from the particle size distribution:-

Effective diameter D10 = .071mm

D30 = .125mm

D50 =.14 mm

D60 =.15 mm

60u

10

C = DUniformity CoefficientD

4.1

So Cu=2.1126

230

c60 10

C =( )

DCoefficient of gradationD D×

4.2

Cc= 1.4671

Percent of silt= 12.22

Percent of fine sand =86.82

Percent of medium sand = .96

4.3.2 Effective Grain Size, D10

The effective grain size of a soil, D10, refers to the diameter of soil particles for which

10 % of particles are finer. D10 is an important value in regulating flow through fills


31


and can significantly influence the permeability of fills. The higher the D10 value, with

the coarser the soil and the better is the drainage characteristics.

The effective grain size is 0.071mm, which is very small signifying that permeability

of fill material will be less. Hazen (1930) related the permeability to the effective grain

size of a soil using equation 2.5

K=Cd102

K= permeability constant, cm/sec

Where d10 = grain size in millimeters for which 10% of the particles pass by weight.

C=constant (≈1)

4.3.3 Coefficient of Uniformity The coefficient of uniformity (Cu) is defined as the ratio D60/D10.

The uniformity coefficient provides a measure of uniformity. Aysen (2002) suggests

that a uniformity coefficient of less then four indicates uniform grading of soil grains.

Whereas, values greater then four indicate a wider assortment of grain sizes. In

general, the uniformity coefficient of the fills was less than four, indicating that the

hydraulic fills were not well graded.

4.3.4 Coefficient of Curvature

The coefficient of curvature (Cc) is defined as the ratio230

60 10( )D

D D×.

Where D60, D30, and D10 are refers of fill particles for which corresponding to 60, 30,

and 10%of particles are finer on the cumulative particle size distribution curve.

For sand, a coefficient of curvature between one and three with the uniformity

coefficient greater then 6, indicates a well-graded soil. Here from the laboratory testing

we have obtained from the laboratory testing Uniformity coefficient Cu = 2.11267,

Coefficient of gradation Cc=1.4671. So this result is also giving information that fill

material is not well graded soil

4.4. Permeability

Falling head permeability tests result is given in Table 4.3.

From table the coefficient of permeability is 4.08×10-3cm/sec which is very less for

hydraulic backfill. Because for hydraulic backfill Coefficient of permeability should be


32


more so that after placement of backfill in open stope in mine, water drainage will be

good. But by this sample, paste backfill will be good because for paste backfill

permeability should minimum so that water will be not drainage.

Table 4.3. Determination of permeability by falling head method

Sample

no.

Length

of water

L (cm)

Height

of soil

∆h (cm)

Area of

tube

A

(sq cm)

Time

t

min

Water

quantity

q (ml)

qL

∆h At

K

(cm/min)

1 11 29 15.21 185 1712 18832 81576.43 0.23

2 11 48 15.21 60 820 9020 43791.26 0.20

3 14.5 44.5 15.21 122 1105 16022.5 82549.57 0.19

4 14.5 24.5 15.21 100 580 8410 37252.99 0.22

5 19.5 37.5 15.21 115 655 12772.5 65572.86 0.19

6 24 42 15.21 65 465 11160 41510.47 0.27

7 28 31 15.21 1020 4010 112280 480791.6 0.23

8 28 21 15.21 145 538 15064 46300.14 0.32

9 32.5 26.5 15.21 262 1055 34287.5 105570.4 0.32

4.5. pH At first determined the weight into a cup. After that added pure water to the sample to

bring the solution to a weight to weight ratio of 1:1. Again stir vigorously for 5

seconds and let stand for 10 minutes .now placed electrodes in the slurry, swirl

carefully and read the pH on pH meter.

Weight of beaker (W1) =102.9gm

Weight of beaker+ mill tailing (W2) = 172.6gm

Weight of beaker +mill tailing +water (W3) = 242.3gm

Result =7.89.

It means mill tailing nature is basic so it is also indicates that in mill tailing sulfate

concentration is negligible.


33


4.6. Unconfined Compression

4.6.1. Compressive Strength and Young’s Modulus The unconfined Uniaxial Compressive Strength (UCS) is calculated as the mean value

of the maximum stresses obtained during the testing of three samples of the same mill

tailing and cement mixture. Table 4.4 summarizes the strengths and moduli obtained

from uniaxial compression testing of fully sealed 54 mm diameter samples for

different cement % and curing period. The secant values of Young’s Modulus are

calculated at the point corresponding to 50% of the compressive strength value (Unal

2000). Different Uniaxial Compressive Strength (KPa) and Young Moduli (MPa) were

obtained for different cement proportions and the curing time.

Fig. 4.5.1 and 4.5.2 show the relationship between UCS and cement %, Young’s

modulus and cement % respectively for different curing period. These figures clearly

show the compressive strength and Young’s modulus of the fill increases with binder

content and curing time as expected. Compressive strength and stiffness are relatively

low for 3% cement in mill tailing, with a notable increase starting to occur for some

mixes with a binder dosage higher than 6%. All samples gradually gained strength and

elastic modulus up to 28 days of curing. These results agree with reports by Belem

(2000). UCS increases nonlinearly with the cement dosage for the all cement mill

tailing composition. Modulus values also follow the same trend, increasing nonlinearly

with binder dosage for all cement mill tailing composition. The rate of increase in UCS

and modulus values is higher in the initial 21 days compared to its increment after 21

days.

High strength values were obtained in the samples containing high amount of cement.

It can also be seen from Fig. 4.5.1 that binders have high strength gain in the

corresponding sample at curing 28 days. Therefore, it can be concluded that longer

curing period also plays an important role for increasing the strength and moduli of

paste backfill. In the present studies it has been noticed that the strength increment is

not as high with curing period as compared to the previous reported works of Lamos

and Clark, 1989. This is may be due to change of temperature in environment.

Fig. 4.5.3 shows the stress–strain curve for different cement composition and for

different curing period.


34


Table 4.4. UCS and Young’s Modulus of Hutti Gold mines paste fill.

UCS(KPa) Curing Time (days) MT:C

3 7 14 21 28

97: 3 173.44 184.6 277.7 299.1 288.6

94: 6 635.25 819.48 929.4 1278.0 1288.2

90:10 1702.8 1863.0 1990.7 2068.8 2075.6

Pulp density

80 %

80:20 5159.3 5727.9 5793.8 5718.0 5867.0

Curing Time (days) Young’s Modulus(MPa)

3 7 14 21 28

97: 3

26.6

27.0 27.3 28.4 91.8

94:06 159.20 187.3 193.7 233.5 237.0

90:10 277.7 331.9 345.1 346.9 467.8

Pulp density 80 %

80:20 654.6 694 705.4 765.1

811.1

0.00

1.00

2.00

3.00

4.00

5.00

6.00

7.00

3% 6% 10% 20%

Cement %

UC

S (M

Pa) for 3 days

for 7 daysfor 14 daysfor 21 daysfor 28 days

Fig. 4.5.1 Effect of Cement and Curing time on Uniaxial Compressive Strength


35


0

100

200

300

400

500

600

700

800

900

3% 6% 10% 20%

Cement %

Youn

g's M

odul

us (M

Pa)

for 3 days for 7 daysfor 14 daysfor 21 daysfor 28 days

Fig. 4.5.2 Effect of Cement and curing time on Young’s Modulus

0

50

100

150

200

250

300

350

0 0.5 1 1.5 2

Strain (%)

UC

S(K

Pa)

curing 3 days

curing 7 days

curing 14 days

curing 21 days

curing 28 days

a

0

200

400

600

800

1000

1200

1400

1600

0 0.5 1 1.5

Strain (%)

Stre

ss (K

Pa)

curing 3 days

curing 7 days

curing 14 days

curing 21 days

curing 28 days

b

0

500

1000

1500

2000

2500

0 0.5 1 1.5

Strain (%)

UC

S (K

Pa)

curing 3 days

curing 7 days

curing 14 days

curing 21 days

curing 28 days

c

0

1000

2000

3000

4000

5000

6000

7000

0 0.5 1 1.5 2

Strain (%)

UC

S (K

Pa)

curing 3 days

curing 7 days

curing 14 days

curing 21 days

cuirng 28 days

d

Fig. 4.5.3 Typical stress- strain curves for paste backfill specimen for different curing period containing (a) 3% cement (b) 6% cement (c) 10% cement (d) 20 % cement.


36


4.6.2 Effect of Pulp Density on UCS Table 4.5 summarizes the strengths and moduli obtained from uniaxial compression

testing of fully sealed 54 mm diameter samples for different pulp density at 28 days

curing period.

Figure 4.5.4 shows the relationship between UCS and different pulp densities for 28

days curing period at 6% cement. Figure 4.5.5 shows the relationship between moduli

and pulp density for 28 days curing period and 6% cement dry weight composition.

When pulp density 83.3% is used then result of UCS of these sample after 28 days

curing in laboratory was 566 KPa, which is 135% more compared with UCS of sample

with pulp density 66.7%. Similarly, Young’s Modulus of pulp density 83.3% sample is

86% more as compared to that of the sample with 66% pulp density (Table 4. 5 and Fig.

4.5.4 and 4.5.5). It can therefore be inferred that the compressive strength and Young’s

moduli of the backfill samples are related to the pulp density. It is also noticed from Fig

4.5.4 that there is no difference in strength values for the samples with pulp densities

between 83.33% and 80 %. This may be due to the required amount of water to react

with cement and develop bonds between tailing materials. The strength of the backfill

decreases as the pulp density decreases mainly because of the subsequent increase in

overall porosity caused by the water-filled voids. On drying these samples air voids are

created which are likely to decrease the strength of samples. On the contrary, the lower

the pulp density ratio the stronger due to greater cement particle interlocking with mill

tailing and less air voids creation.

Fig. 4.5.6 also shows the stress –strain curve at different pulp density for 6% cement.

Table 4.5 Effect of pulp density on UCS and Young’s Modulus.

Pulp density (%)

Composition

(MT:C:W)

UCS

(KPa)

Young’s Modulus

(MPa)

83.3 94:6:20 566.3 157.1

80.0 94:6:25 503.3 127.9

76.9 94:6:30 369.4 128.5

74.0 94:6:35 269.4 121.0

71.4 94:6:40 242.9 86.9

66.7 94:6:50 240.8 84.3


37


566.3503.3

369.4

269.4 242.9 240.8

0

100

200

300

400

500

600

83.3 80 77 74 71.4 66.7

Pulp density (%)

UCS

(KPa

)

Fig.4.5.4 Effect of Pulp density on uniaxial compressive strength

157.1

127.9 128.5121

86.9 84.3

0

20

40

60

80

100

120

140

160

180

83.3 80 77 74 71.4 66.7

Pulp density (%)

Youn

g's

Mod

ulus

(MPa

)

Fig.4.5.5 Effect of pulp density on Young’s Modulus

0

100

200

300

400

500

600

700

0 0.5 1 1.5

Strain (%)

Stre

ss (K

Pa)

PD 83.3 %PD 80%PD 77 %PD 74 %PD 71.4%PD 66.7 %

Fig. 4.5.6 Typical Stress- Strain curves at different pulp density for 6% cement

(dry weight %)


38


4.6.3 Effect of Porosity on Uniaxial Compressive Strength Porosity has decreased with addition of cement in mill tailing due to fineness of cement.

So when we mix cement in mill tailing then void ratio of mill tailing decreased. So

higher uniaxial compressive strength has found in lower porosity due to greater particle

interlocking and the presence of more cement is available per unit volume of backfill.

The effect of cement addition is given in table 4.6. And effect of porosity on uniaxial

compressive strength is given in fig. 4.5.7.

Table 4.6: Effect of Cement addition on Porosity and UCS

Cement content in

Mill tailing (%)

Porosity (n %) UCS(KPa)

3 60 173.44

6 59.18367 635.25

10 58.69565 1702.8

20 55.5556 5159.3

173.44635.25

1702.8

5159.3

0

1000

2000

3000

4000

5000

6000

55 56 57 58 59 60 61

Porosity (n%)

UCS

(KPa)

Fig. 4.5.7 Effect of Porosity on UCS for (after curing 3 days)

4.7. Shear Strength Parameters

Normal stresses required for testing were estimated by dividing the applied

load by the area of the shear box. Peak shear strength was determined from plots of

shear stress versus shear strain. Internal friction angle was obtained using a linear best-

fit line from the plot of peak shear strength versus normal stress. The residual friction

angle was obtained using a similar best-fit line. Fig. 4.6.1 shows the variation of shear

stress with shear strain. Table 4.7 and fig 4.6.2 show the shear strength with normal

stresses which gives internal friction angle 300 and cohesion. 17.4 KPa.


39


0

10

20

30

40

50

60

70

0 1 2 3 4 5 6 7 8 9 10

Shear Strain (%)

Shea

r Str

ess

(KPa

)

normal stress 19.6 Kpanormal stress 35.28 KPanormal stress 47 KPanormal stress 58.8 KPanormal stress 70.56 KPa

Fig. 4.6.1 shear strain vs. shear stress curve from direct shear test

Table 4.7. Variation of shear strength with normal strength

Normal Stress (KPa)

Shear Strength (KPa)

19.6 31.85

35.28 35.53

47.00 40.15

58.80 50.63

70.56 60.57

y = 0.5697x + 17.4R2 = 0.928

0

10

20

30

40

50

60

70

0 20 40 60

Normal stress (KPa)

Shear S

tren

th (K

Pa)

80

Fig.4.6.2 Normal stress vs. Shear strength from direct shear test


40


CHAPTER 5

BACKFILL STRENGTH ENHANCEMENT

5.1. Introduction The evolution of backfill technology is closely related to the advancement of modern

mining methods and the development of new backfill technology continues to evolve.

Backfill operations employ large quantities of cement and experience large costs

associated both with backfilling practice and cement use. Opportunities exist for

reducing backfill energy costs and enhancing the application of alternative cementing

agents in backfill. One of the techniques is the application of superplasticizer

[Stuart.1980] which is equally effective both in terms of enhancing the strength as well

as in the cost. This project paper presents the results of only two design stages

associated with the superplasticizer: Strength development and flow performance. The

advancement of this technology has reached a stage that should promote industrial

implementation within the Indian mining industry. A number of Indian mining

operations exhibit potential for plant optimization through possible binder source

replacement, and might thus realize associated reductions in backfill and energy costs.

Utilization of superplasticizer in mine backfill would create a new market for

manufacturing of this, so that its cost will be reduced.

5.2. Effect of Super Plasticizer on Compressive Strength and Young’s Modulus For the cemented paste backfill with superplasticizer specimens, the same procedure

was performed, but differed by adding the superplasticizers after 5 minutes and mixing

the superplasticizer with paste backfill for extra 2 minutes in mixing bucket.

The results of all the UCS tests due to variation of superplasticizer are summarized in

table 5.1 and fig. 5.1 and 5.2 for the different % of composition after 28 and 90 days of

curing. A typical stress- strain curve is shown in the fig.5.3 and 5.4 for the different %

of composition after 28 and 90 days of curing.

The unconfined compressive strength was calculated as the arithmetic mean of the

maximum stresses obtained during the testing of three samples of the same paste


41


mixture. Secant values of Young’s modulus are calculated at the point corresponding to

50% of the compressive strength value

Table 5.1. Effect of SP on Compressive strength and Young’s Modulus for 28 and 90 days curing

for 28 days curing

Composition UCS(KPa)Young's Modulus(MPa)

94:6:.2 654.26 214.2 96:4:.2 545.91 165.36 94:6:0 313.64 124.84 97:3:.3 130.16 150.90

for 90 days curing 94:6:.2 938.9 234.0 96:4:.2 892.1 198.6 94:6:0 586.1 170.2 97:3:.3 331.87 97.3

654.26545.91

313.64

130.16

938.9 892.1

586.1

331.87

0100200300400500600700800900

1000

94:6:.2 96:4:.2 94:6:.0 97:3:.3

Composition (MT:C:SP)

UCS

(KP

a)

for 28 daysfor 90 days

Fig.5.1 Effect of Superplasticizer on UCS for different curing days


42


214.2

165.36

124.84150.9

234

198.6170.2

97.3

0

50

100

150

200

250

94:6:.2 96:4:.2 94:6:.0 97:3:.3

Composition (MT:C:SP)

You

ng's

Mod

ulus

(MP

a)

for 28 daysfor 90 days

Fig.5.2 Effect of superplasticizer on Young’s Modulus for different curing days

0100200300400

500600700800

0 0.2 0.4 0.6 0.8 1

Strain (%)

Stre

ss (K

Pa)

.2% SP mixed withMT:C (94:6) .2 %SP mixed withMT:C (96:4) no SP mixed withMT:C (96:4) .3% SP mixed withMT:C (97:3)

Fig.5.3 Typical Stress- Strain curve after 28 days of curing for the different % of composition


43


0

200

400

600

800

1000

1200

0 0.2 0.4 0.6 0.8 1 1.2Strain (%)

Stre

ss (K

Pa)

.2% SP mixed withMT:C (94:6) .2% SP mixed withMT:C (96:4) no SP mixed withMT:C (96:4) .3% SP mixed withMT:C (97:3)

Fig. 5.4 Typical Stress- Strain curve after 90 days of curing for the different % of composition

Fig.5.1 and 5.2 show the variation in the UCS and Young’s Moduli with the variation of

composition of paste backfill with superplasticizer for 28 and 90 days curing. Fig. 5.1

shows the maximum compressive strength 654.26KPa (just double) of the composition

MT:C:SP containing 94:6:.2 ratios respectively as compression to compressive strength

313.64 KPa of composition MT: C: SP containing 94:6:0 ratios (control binder)

respectively for 28 days curing. Compressive strength of another binder in which MT:

C: SP containing 96:4:.2 ratios are also 74% more strength as compression to that

control binder. But effect of superplasticizer is not good in binder which contains MT:

C: SP containing 97:3:.3 ratios. Compressive strength 130.16 KPa of this binder is less

than half value of compressive strength of control binder. Fig. 5.2 shows Young’s

modulus 214.19 MPa of binder which contains MT: C: SP containing 94:6:.2 ratios is

also 70 % more than that of the Young’s modulus of control binder. This type of

increment in compressive strength and stiffness has happened due to renders a lower

porosity hardened material and increased the rate of cement hydration in well dispersed

cement so that between cement –mill tailing better particle packing and denser

structure upon hardening in pastes contains admixture superplasticizer.


44


Fig. 5.1 and 5.2 clearly show the variation of curing time on its strength and moduli.

Increment on strength due to curing varies from 50- 100% for different composition.

This has happened may be due to long term hydration between cement and mill tailing.

Fig. 5.3 and 5.4 produced not cleared stress-strain relationship for 28 and 90 days

curing.

Superplasticizer affects the unconfined compressive strength with curing. The cement

paste backfill mixture MT: C: SP containing 94:6:.2 developed the highest unconfined

compressive strength over a 90 days curing period and showed the maximum stiffness

development as compared to with other those of paste backfill specimens without

admixture.

But the cement paste backfill mixture MT: C: SP containing 96:4:.2 also developed the

required unconfined compressive strength over a 90 days curing period and showed

the maximum stiffness development as compared to with other those of paste backfill

specimens without admixture. So for economical purpose this composition is also best.


45


CHAPTER 6

RHEOLOGICAL TEST

6.1. Laboratory Preparations The materials used for characterizing the rheological properties were non-plastic mine

tailings from a Hutti gold mine.

The two types of homogenous solid-liquid mixtures characterized in this study were

cemented paste backfill with superplasticizer and other without superplasticizer

cemented paste backfill, in both case pulp densities was used 77%. For the cemented

paste backfill without superplasticizer specimens were prepared by adding 6% cement

and 23% tap water to mill tailing, the water was mixed for 5 minutes in the mixing

bucket. For the cemented paste backfill with superplasticizer specimens, the same

procedure was performed, but differed by adding the superplasticizer after 5 minutes

and mixing the superplasticizer with paste backfill for extra 2 minutes.

6.2. Experimental Procedure Cylindrical mould was used for determination of slump value. There is no required

standard for the cylinder test. Cylinder was made by PVS with the length 115 mm and

diameter 102 mm. The both side of the cylinder was opened so that slumped material

is 100% consistent during lifting. And filling with sample time used 1 strong smooth

steal plate on top of cylinder. The cylinder was filled with sample, and the cylinder

lifted slowly and evenly. The change in height between the cylinder and deformed

material was measured (fig.6.1). The midpoint of the slumped material was taken as

the representative height. Heights were measured with a scale. Density and

concentration were measured at the time of testing. Average value was finding of three

tests for each. Cylindrical slump test (a) with superplasticizer (b) without

superplasticizer


46


Fig. 6.1 Slump test of backfill with Superplasticizer

6.3 Calculation Length of PVC Cylinder = 115 mm

Diameter of PVC Cylinder = 102 mm.

So volume of Cylinder = 9.4× 10-4 m3

Weight of material used in cylinder =2 kg

So slurry density of material = 2127.66 kg/m3

For without superplasticizer Height of undeformed region = (115-20)/2 = 47.5

Dimensionless height of undeformed region, ' 00

hhH

=

h0' =47.5/115 = .413

The dimensionless yield stress of backfills, '0

1 h '2yτ =

τ'y =.413/2 = .206

Yield stress of backfills τy = τ'y× ρgH

τy= .206×2127×9.8×.115 = 493 Pa

Dimensionless slump )]2ln(1[21s yy τ′−τ′−=′

s' = .2226

So slump height = .2226× 115 = 25.6 mm


47


For with Superplasticizer

Height of undeformed region = (115-50)/2 = 32.5

Dimensionless height of undeformed region, ' 00

hhH

=

h0' =32.5/115 = .282

The dimensionless yield stress of backfills, '0

1 h '2yτ =

τ'y =.282/2 = .141

Yield stress of backfills τy = τ'y× ρgH

τy= .141×2127×9.8×.115 = 338 Pa

Difference of yield stress between both = 155 Pa

Dimensionless slump )]2ln(1[21s yy τ′−τ′−=′

s' = .36

So slump height s = .36×115 = 41.51 mm

Difference of yield stress between both = 155 Pa

Difference of slump height between both =41.51-25.6 =15.91 mm

6.4 Result

Yield stress of backfills without superplasticizer = 493 Pa

Slump height = 25.6 mm

Yield stress of backfills with superplasticizer =338 Pa

Slump height = 41.51 mm

The results of the slump tests performed with 6% (dry weight %) cements with .2%

superplasticizer and without superplasticizer (fig. 6.2). It has seen that there is very

much difference (155 Pa) on the yield value and also much difference between slump

height(15.91mm) between both while water 23% in solution present in both condition.

So here fluidity increased with superplasticizer.


48


493

338

0

100

200

300

400

500

600

without SP with SP(.2%)

Yie

ld S

tres

s (P

a)

Fig.6.2 Effect of Superplasticizer on Yield Stress

6.5 Setting Time

6.5.1 Procedure Setting time was determined by Vicat needle test (penetration test). Specimen For the

vicat needle test were cylindrical cup 70 mm in diameter and 40 mm high. After being

filling with paste (pulp density 80%). The standard test method, the Vicat needle test,

was used to determine the initial and final setting times hydraulic cement. The initial

setting is the determined for the needle to reach a penetration depth 5mm in standard

Vicant apparatus. The final setting takes place when the needle does not visibly

penetrate into the paste i.e., the specimen has a solid structure.

6.5.2 Result Result is given in table 6.1.

Table 6.1.Vicat needle test result

Setting time(min) Binder

content

Additive

Water(%) of total weight

Initial Final

94:6 MT:C none 23 65 125

94:6 MT:C SP.2% of dry

weight 23% of total weight 45 120


49


Table 6.1 shows the initial and final times of setting for paste in which one is without

superplasticizer and other is with superplasticizer. The data indicates that initial time

setting of paste with superplasticizer is less as compared to without superplasticizer

paste. And final time setting of both pastes is about same while superplasticizer paste

is wet as comparison to without paste. So we can say for same slump value time

setting will be reduced in superplasticizer paste than without superplasticizer paste.

6.6. Flowability Test For flow behavior test, one galvanized iron sheet 120 cm length was used at inclination

200 degree as shown in figures 6.3-6.7.

Figures 6.3-6.7 show the flow characteristics of backfill material. The result of the test

performed with 4 different compositions. In first experiment for flow test, .2 %

superplasticizer was used in MT: C contains 94:6 ratios binder. In 2nd experiment no

superplasticizer was used for same combination. In 3rd experiment, .2%

superplasticizer was used in MT:C contains 96:4 ratios respectively. In 4th experiment,

.3% superplasticizer was used in MT: C contains 97:3 ratios respectively. It is seen

that there is significant difference on the fluidity of different composition. At .2%

superplasticizer in Mill Tailing –Cement (94:6 ratios) binder, the fluidity increased

compared to the other composition. And in other composition some part of paste has

flowed and some part has not flowed. Higher the fluidity in first case was observed

due to electrostatic repulsion between particles, causing dispersion (Nkinamubanzi and

Aitcin 2004). In 3rd and 4th experiment an insufficient amount of cement may be

available to react with main hydration (i.e. calcium silicate hydrates or C-S-H) to

produce effective dispersion at later stage. Fine particle is also important role played

with superplasticizer for fluidity purpose.

The rheological behaviour of two paste backfills characterized in this study was yield-

pseudoplastic. The superplasticizer controls not only the rheological behaviour of paste

backfill, but also their yield stress. Yield stress measurements in slump test method

show reliable results for superplasticizer as comparison to non superplasticizer paste

backfills. So Based on the results of this research, we can conclude that the use of

superplasticizer in backfill material will be economical because this will not increase

the strength but also aids in the rheological characteristics of paste backfill material.


50


Fig.6.3 Flow characteristics of mill tailing when .2 % superplasticizer mixed with MT: Cement (94: 6)

Fig.6.4 Flow characteristics of mill tailing When no superplasticizer used in

MT: Cement (94: 6)


51


Fig.6.5 Flow characteristics of mill tailing when .2 % superplasticizer mixed with MT: Cement (96: 4)

Fig.6.6 Flow characteristics of mill tailing when .3 % Superplasticizer mixed with MT: Cement (97: 3)


52


CHAPTER 7

CONCLUSIONS

An effort is made in the present study to generate value on physical, chemical,

mechanical and rheological characteristics of Hutti gold mill tailing samples.

Mechanical and rheological characteristics of mill tailings were found with different

percentage of cement, water and superplasticizer. Ninety different cemented backfill

materials were tested to determine the influence of their composition on the

unconfined compressive strength in laboratory.

7.1 Conclusions The studies lead to the following conclusions which may be of significance for the

further research and for the practice of mine backfill using cemented paste backfill

method.

• Predominant oxides found in the Hutti mill tailing samples are SiO2, Fe2O3, Al2O3,

CaO, Na2O, MgO, SO3, and TiO2. The sums of these oxides were above 90%. The

presence of CaO at 9 % in the mill tailing samples indicates the strong pozzolanic

characteristic of mill tailings.

• Generally mill tailings contain sulfate concentration. pH was found to be at 7.89 in

this sample and by SEM very less sulfur concentration has been observed. So it is

also good for strength purpose. Because sulfate concentration decreased strength

after 90 days. Due to this cause dilution problem will be decreased.

• Particle size distribution show that the Percent of fine sand is 86.82%, for paste

backfill purpose minimum15% of below size 20μm mill tailing will be required

will be required.

• Coefficient of permeability of mill tailing is 4.08×10-3cm/sec, which is very less

and after cement, addition its value again will be decreased. So it is not good for

drainage in hydraulic backfill purpose without any flocculent. This is good for

paste backfill purpose. After addition of cement and fine particle permeability will

be decreased.


53


• One hundred twelve different cemented backfill materials were tested to determine

the influence of their composition on the unconfined compressive strength in

laboratory. 60 samples were tested with different percentage of cement at 80%

pulp density. 18 samples were tested with different pulp density. And 24 samples

were tested with different percentage of binder with superplasticizer.

• The material composition strongly influenced the strength of cemented backfill.

Pulp density is a critical determining factor in the strength of cemented backfill.

Increase in its value significantly increased the backfill strength.

• Increased cement content increased the backfill strength.

• Superplasticizer also play good impact for increment on its strength with cement,

• Slump heights obtained from the slump cylinder experiments is observed to

increase and decrease yield stress with mixing of superplasticizer in binder. So its

flowability increased with mixing of superplasticizer. Setting time has also not

increased with superplasticizer. Due to these good properties of superplasticizer, it

may play a vital role in paste backfill.

So the use of paste backfill in place of hydraulic backfills will be correct choice for

backfill operation. This shall not only enhance the performance of the backfill as a

ground support system but also likely to reduce the dilution of muck, and thus may

result in the full recovery of ore.

7.2 Scope for further Studies

• Numerical modeling is required for determination of accurate strength

requirement of backfill for stability in stope.

• Transportation of paste backfill could not be studied. Transportation rate and

quantity of discharge from surface to stope should be conducted both

numerically and experimentally to model and verify long term characteristics.


54


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APPENDIX

BACKFILL DESIGN

The aim of backfill design is to develop a procedure for calculating the fill strength

requirement for a different height, strike length and width of fill exposure.

Various authors have published backfill design techniques and these are briefly

reviewed in the following section. The techniques discussed are:

1. Freestanding Vertical Faces

2. Vertical Slopes

3. Limit Equilibrium Wedge,

4. Arching.

A1. Free Standing Vertical Face Mitchell et al (1982) note that the largest shear stresses in backfill are caused by self-

weight. The cemented backfill can be designed as a freestanding vertical face. A

freestanding wall where the unconfined compressive strength required at any depth in

the fill is given as

σz≥ ρg z A.1

Where

σz =unconfined compressive strength (kPa)

ρg =bulk density of backfill (KN/m3)

z =depth of backfill from the surface (m)

A2. Vertical Slopes

Mitchell et al (1982) discuss the vertical slope method where φ= 0 for a constant

strength fill

=2zgHρσ A.2

Where: H = overall height of exposed fill (m)


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A3. Limit equilibrium wedge by Mitchell

Fig.A.1. Confined block mechanism (Mitchell et al. 1982)

Mitchell et al. (1982) modeled the failure of a single exposure 3-D fill mass as shown

in Fig.A.1. A shear plane for sliding failure was defined within the ‘block’. By

assuming that there exists some shear resistance between the fill and stope walls due to

the fill cohesion, the design uniaxial compressive strength required to maintain

stability with safety factor of F can be evaluated by using following relationship

(Mitchell et al. 1982):

A.3

Where γ is the Fill bulk unit weight (kN/m3), α is the angle of failure plane from

horizontal (= 45 + /2), is the friction angle of fill (Degree), c is the cement bond

strength ( cohesive strength) of fill (kPa), l is the strike length of block (m), w is the

width of block (m), h is the height of the block (m), h* is the height of block from top

to the centroid of the triangular section of the sliding wedge (m), F is the Factor of

safety.

In the long term, the compressive strength of the fill material is mainly due to binding

agents and strength contributed by friction can be neglected (i.e. = 0). For a

frictionless material, cohesion is considered as half of the uniaxial compressive

strength (UCS) (i.e. c = UCS/2). Therefore, Eq. (A.3) can be modified as:


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A.4

The stability of a free standing backfill can also be determined when height of the

stope to be much greater than width. The required strength (UCS) of frictionless fill is

given by;

A.5

A4. Extended Arching - Marston theory by Pirapakaran (2006)

The analytical solution presented arching theory by Marston (1930), Terzaghi (1943)

and Aubertin et al. (2003) are for a 2-dimensional stope where the fill is subjected to

plane strain loading. In reality, mine stopes are rarely 2-dimensional, and therefore it is

quite useful to extend these theories to three dimensions.

An attempt is made here to extend the above theories and develop expressions for

vertical and horizontal stresses within a 3-dimensional mine fill stope

A schematic diagram of a 3-dimensional stope is shown in Fig. A.2.(a) With the

dimensions. Figure A.2 (b) shows the free body diagram of the forces acting on an

infinitesimal horizontal layer within a vertical stope, where h is the backfill height, w

the stope width, dh the thickness of the layer element, W is the weight of the backfill

above the layer element. dC is the lateral compressive force, dS is the shearing force at

the fill- rock interface and V and V+dV are vertical forces at the position h and h+dh

respectively.

Weight of the element

A.6

Compressive forces acting on the vertical faces

A.7

The shear force S is defined as:

A.8

Where δ is the friction angle between the backfill and the wall it cannot be greater than

(may be assumed between 1/3 to 2/3)


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For the equilibrium of the small element,

A.9

Here, it is assumed σv uniformly distributed over

Fig. A.2. Schematic diagram of a 3-dimensional stope (a) and the free body diagram With forces (b) the entire width w (Pirapakaran 2006)

A.10

Earth pressure coefficient from soil mechanics theories is defined as;

A.11

The following relationships can be found out from equations (A.6)–(A.11);

A.12


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A.13

By integrating Eq. (A.13), vertical and horizontal stresses which are acting within the

stope can be found out as follows;

A.14

A.15

So for the strength requirement purpose I assume the following condition

1. The fill material is non cohesive i.e. c=0,

2. The friction angle between the backfill and wall δ = 1/3 ,

3. Rock’s stiffness is stiffer than backfill material so that rock would be in rest

condition i.e. K=Ko =1-sin

4. Factor of Safety =1.5

1 exp 2 tan2 tandesign

w l l wUCS Kh FSK l w lwγ δ

δ⎡ ⎤⎧ + ⎫⎛ ⎞ ⎛ ⎞= − −⎜ ⎟ ⎜ ⎟⎨⎢ ⎥+⎝ ⎠ ⎝ ⎠⎩ ⎭⎣ ⎦

×⎬ A.16


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0

200400

600800

1000

12001400

1600

0 15 30 45 60 75 90 105

Height (m)

UCSr

eq (K

Pa)

l=60 m & w =5 ml=60 m & w=10 ml=60 m & w=20 ml =80 m & w =20 ml= 100m & w=10 m

Fig.A.3. UCSreq vs. height of fill for different width and length in backfill

Fig.A.3 shows strength requirement in backfill material increases with height, width

and strike length of fill. Here figure shows Impact of width, height is more than length

on strength.

From fig. A.3 we can say that uniaxial strength for narrow mining width 5

m, height 45 m and length 60 m (e.g. Hutti gold Mines) will be required approximate

515 KPa for stability of backfill material in stopes. We know that temperature also

play important role in backfill strength. Strength increases with temperature up to

control condition.. So we can say that in Hutti gold mines strength of backfill will be

increased than these laboratory values. Because 3% cement was given 288 KPa after

28 days curing in paste backfill composition and UCS of 6% cement was given 1288

KPa which is more than required value. So for safety and economical purpose we can

use cement in mill tailing between 3 to 6% in paste backfill.


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Documents

Influence of Cemented Mill Tailing as Backfill Material in Open Stopes