2. Page | i
Executive Summary
This report presents the methods used to give a detailed specification for a DR Flotation Machine,
contained in section four of Situsβ proposed potash processing plant in Boulby, Teesside. The report is
one of a total of six reports written by Situs to aid the development of the project, which aims to
commission a potash processing plant capable of producing one million tonnes of saleable product
annually.
The process has been broken into six sections which have all been individually reassessed and are
described in the report. Section four contains the froth flotation stage of the process, however it also
contains a salt separation sequence included as an added value stream. The report gives a basic
specification for all of the equipment found in section four, along with a detailed specification for a
DR Flotation machine.
The DR Flotation machine is capable of separating the desired KCl particles from the rest of the
material by utilising their differences in surface chemistry. The introduction of flotation reagents
causes the surfaces of the KCl particles to become hydrophobic, thus causing the particles to float out
when they come into contact with air bubbles.
Potash ores contain water-soluble salts, meaning flotation has to be carried out in a saturated brine
solution, which can be highly corrosive. Therefore, the majority of the equipment in section four will
be constructed from SS-316.
The report also demonstrates the control strategy chosen by Situs and how it will be implemented
across section four in the form of a P&ID. A HAZOP and other hazard studies have been conducted
over recent weeks and used the P&ID to identify potential hazards that could be encountered during
process operation.
An economic overview of section four has also been included in the report, and will be a key part of
the final group report. The total capital cost of section four was estimated to be Β£9,859,000. Situs have
also estimated the annual operating cost of section four to be Β£3,670,000. The potential revenue will
be available to view in the final group report.
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1. Introduction
1.1. Project Overview
A variety of potassium-bearing minerals are often referred to as potash, however only those that are
soluble in water are of major commercial interest to Situs. Potassium chloride, also known as sylvine,
is the most significant source of potash globally and is currently mined at Boulby, Teesside, situated
in the North York Moors National Park.
This project aims to design a new potash processing plant capable of producing one million tonnes of
saleable product, from three million tonnes of ore, annually and will be situated at the site of the
existing mine in Boulby. The design of the new plant takes into account sustainability, environmental
issues and profitability.
1.2. Aims & Objectives
The report aims to clearly describe and define the technical design of section four of the process,
proposed in the previous report. Section four consists of two flotation stages, along with the salt
separation process. A cell found in the rougher flotation stage will be the subject of the detailed design,
whilst specifications will also be given for the remaining pieces of equipment found in section four.
The report will also include the relevant control implemented across section four and the control
philosophy in place for the remainder of the plant. Following the technical design and area
specification, there will be an economic evaluation containing both capital and operating costs for the
section, along with a thorough conclusion.
The report will recognise the essential aspects for specifying and designing a cell found in the rougher
flotation stage of the process, along with the surrounding area of the plant, however to do this
effectively the key objectives must first be made clear. The scope of the report will include:
- A complete and comprehensive overview of the potash refining process
- A summary of the health, safety and environmental considerations
- Company polices, legislations & standards
- A revised and more detailed mass balance covering section four
- Specification of equipment used for the key unit operations encountered in section four
- Detailed chemical engineering design of a cell found in the rougher flotation stage
- Description of control philosophy employed over the entire plant, with focus on section four
- An economic review of the section under consideration
- A summary of the reports key findings
1.3. Process Overview
The process will run on a continuous Twenty-Four/Seven schedule and have planned shutdown
periods for seven days in May and fourteen days in August. The following overview has split the
process up into six sections, depending on the general purpose of the unit operations found in that
area.
1.3.1. Section 1
The aim of section one is to reduce to size of the particles of ore to such an extent that the desired
product can be liberated from them, figure 1.1 shows a process flow diagram of the section under
consideration. The size specifications that this design will be working from are those given by the client
β both for the size distribution of the ore being fed into the plant and the required size distribution of
the final products. This reduction is from an average particle diameter of around 1500 mm to an
average particle diameter of around 0.5 mm, with a maximum particle size of roughly 2.5mm. The
7. Page | 2
crushing process used follows the common model of: three stages of crushing followed by one stage
of grinding. This method is commonly used in similar mineral processing plants (Metso, 2008).
The initial crushing unit is a jaw crusher and it will operate on a batch basis (Metso, 2008). The jaw
crusher will feed a storage vessel, large enough to contain 24 hours worth of material that will allow
the rest of the process to run on a continuous basis (Metso, 2008). Cone crushers carry out secondary
and tertiary crushing. The purpose of these is to reduce the size of the particles of ore to a size suitable
for grinding. Between each of the crushing stages will be a vibrating screen, able to filter out the
oversize that must be recycled to the previous crushing stage, and the undersize that can be scalped
and skip the next crushing phase to prevent clogging the equipment.
The grinding stage of the process, carried out using a ball mill, liberates the desired potassium minerals
from the ore. Before entering the ball mill, the solid particles are mixed with brine to increase
efficiency (Wills & Napier-Munn, 1985). Following the ball mill, the particles are passed through an
attrition scrubber to remove the insolubles from the surface of the desired particles, before moving
onto section 2.
1.3.2. Section 2
The feed from section one contains a large amount of insolubles which the attrition scrubber, prior to
the primary hydrocyclone bank, has liberated for separation. Brine is added to the feed into the
hydrocyclones to reduce the solid mass percentage. The primary hydrocyclones are designed such
that they cut at 50 Β΅m.
The underflow from the primary hydrocyclone bank is fed onto an attrition scrubber. By liberating
further insolubles from the KCl, the adjacent flotation stage can operate at optimal efficiency whilst
the remaining slimes can be recycled. The outlet from the attrition scrubber is injected with more
brine before being fed into a secondary hydrocyclone bank, cutting at 100 Β΅m. The overflow is recycled
back to the primary hydrocyclone bank and the underflow continues onto section four for flotation.
The overflow from the primary hydrocyclone bank contains a large amount of brine that must be
recycled before further processing. The flocculant is added and the fines and slimes are thickened
Figure 1.1: Process flow diagram of section one.
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from 4% solids to 15% in a continuous thickener. The brine in the overflow is recovered and the
underflow continues on to a tertiary hydrocyclone bank. The tertiary cyclone bank is designed to rid
the stream of all slimes and little KCl. Clays typically range from 2-8 Β΅m and most of the KCl is generally
much larger, having the hydrocyclone cut at 10 Β΅m ensures this is possible. The overflow is then taken
to backfill whilst the underflow continues to fines flotation in the Jameson cell. In the Jameson cell the
KCl fines are selectively floated from the insoluble clays and any NaCl. The Jameson Cell increases the
mass % of KCl from 30% in the feed to approximately 75% in the concentrate. The concentrate then
continues onto section three to produce the damp cake product, whilst the tails report to a second
thickener for brine recovery. Figure 1.2 shows a process flow diagram of section two.
1.3.3. Section 3
Section three is responsible for the production of the damp cake product and is illustrated by the PFD
in figure 1.3. The feed to section three from the Jameson Cell enters a balance tank allowing the
remainder of the section to run on a continuous basis. The next stage of the process is heating the
materials to 90o
C inside a boiler before feeding it to a mixing tank where it will be combined with
enough fresh water to bring the KCl concentration to 25% by mass. After thorough mixing the solution
passed through a candle filter, removing all the insoluble clays, 90% of the NaCl and 10% of the water.
Figure 1.2: Process flow diagram of section two.
Figure 1.3: Process flow diagram of section three.
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The residue from the filter is sent to backfill. After filtration the KCl solution is fed to a centrifuge to
reduce the water content to 10% and then continues to the second balance tank. The purpose of the
second balance tank is to store the incoming KCl concentrate whilst also feeding the dryer on a batch
basis. After drying to remove half of the remaining moisture the damp cake KCl is kept in storage.
1.3.4. Section 4
The first unit found in section four is the conditioner, as can be seen in figure 1.4. The Process flow
diagram shown in figure 1.4 contains section four and five as they are largely dependent on one
another, making it more logical for them to be illustrated as a pair. In the original process proposed in
the concept report, the fine particles were to be conditioned separately to the coarse particles, to
improve the efficiency of the reagent usage. However, as many of the fine particles are removed and
processed separately in the damp cake section, the two conditioners have now been replaced by a
single one.
After conditioning, flotation takes place using three consecutive DR flotation cell systems before finally
a column is used for the scavenger system, however only two of these are found in section 4; the
rougher and cleaner. The recleaner and scavenger systems are found in section 5. The first stage is the
rougher flotation, which is designed to float as much as KCL as possible (Perucca, 2001). The rougher
concentrate is passed through a 0.84 mm screen where the oversize particles are sent to the
centrifuge in section 5 and the undersize particles are passed on to a cleaner flotation system designed
to generate a purer product, whilst the rougher flotation tails are passed through a 1.41 mm sieve
bend screen (Perucca, 2001). The cleaner concentrate is passed onto the recleaner stage found in
section 5, whereas the tails are mixed with the rougher tails prior to entering the sieve bend screen.
The flotation tails from the recleaner and the scavenger cells are then combined with the underflow
from the sieve bend screen. The first stage is to pass the stream through a screen to separate coarse
NaCl from fine clays. This simple method can be effectively utilised due to the large size difference
between clay fines and salt crystals. Salt crystals are commonly sized between 0.1 and 2mm whilst
clay fines can be as small as 10Β΅m (Muttiah, 2002).
Figure 1.4: Process flow diagram of section four & five combined.
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The coarse NaCl overflow will then be dried. The underflow from the screen will contain salt particles
of less than 500Β΅m and all the clay fines. In the concept report it was proposed that a hydrocyclone
would separate the NaCl fines from the insoluble fines. However, this has now been removed from
the process as the value added from the isolation of the fine NaCl particles was not worth the energy
spent retrieving them. The underflow stream from the screen is then run through a dewatering
hydrocyclone. The brine from which is recycled back into the process and the fines taken to backfill.
1.3.5. Section 5
This section deals with the processing of the flotation concentrates, as well as the second half of the
flotation circuit, discussed in the previous section. Tails from the rougher flotation cells will primarily
contain NaCl particles and the larger KCl particles which will not have been floated due to their slower
kinetics. This feeds through a sieve bend screen with an overflow to be sent towards the scavenger
flotation. This contains large KCl particles as well as a proportion of NaCl. Prior to entering the
scavenger flotation column, the material is regrinded and reconditioned.
The concentrate from the cleaner flotation cells shown in section 4 is fed into the recleaner flotation
cells in order to help improve the purity of product. The concentrate from the recleaner cell continues
onto the centrifuges to be dewatered whilst the tailings are sent to the salt separation unit shown in
section 4. The centrifuges reduce the liquid content to 8% or lower forming a damp cake and then
send this onto the lump impactor. The cake is transported on conveyor belts into a fluidised bed dryer.
The dryer is supplied by very hot air from a gas heated furnace. This dries the cake down to 0.1%
moisture. The hot gases released by the dryer contain fine KCl particles which mustnβt enter the
atmosphere as they can be hazardous to personnel. As a result, the exhaust gases are passed through
a set of cyclones and an electrostatic filter to remove all entrained particles.
1.3.6. Section 6
This section focuses on final stages involved with the production and storage of the coarse, standard
and white product. A process flow diagram of section six can be seen in figure 1.5.
Figure 1.5: Process flow diagram of section six.
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A holding tank is needed to contain the dried KCl to ensure a continuous feed to the double-deck
screens. Three types of particles are separated out through the double deck screens: coarse, standard
and fine particles. The overflow from the top screen is contained in the coarse product storage and
the overflow from the bottom screen is contained in the standard storage. The fine particles in the
underflow are below 50ΞΌm and are passed onto a recrystallization circuit to form white product (Wist,
et al., 2009).
The fines are mixed with enough hot water to dissolve all of the KCl. This solution then enters a
centrifuge which removes the majority of the impurities, before being pumped into a draft tube baffle
crystallizer. The feed is directed upward into the draft tube where adiabatic evaporative cooling takes
place (Batten, et al., 2000). The KCl that crystallizes out of the solution is removed from the water
using another centrifuge. Following this the KCl is dried in a fluidised bed dryer which produces the
white product.
1.4. Health, Safety & Environmental Considerations
The potential impact a chemical plant can have on its surrounding area must always be taken into
consideration. A major environmental concern of any mining process is surface deformation, and can
be avoided by the process of backfilling. The process of backfilling is particularly advantageous as the
waste produced is the same as what was taken from the ground initially, due to the low number of
chemical reactions in the process. The proposed design aims to backfill all of the waste produced to
minimise the adverse effects it could have on the environment.
Noise is a problem often encountered by mineral processing plants; therefore ear protection will be
required in certain areas of the plant, along with the rubber lining of some machinery found in the size
reduction section of the process to minimise the noise produced. Further information on the consent
levels adhered to on site can be found in Hazard Study 1.
The plant is operated on a wet basis where possible, to minimise the release of dust into the
atmosphere. Dust pollution cannot be prevented in all cases; therefore dust masks will be an essential
part of the PPE on site. A comprehensive list of the PPE required on site can be found in Hazard Study
1.
In recent weeks further hazard studies have been carried out on the new processing plant that has
been proposed. Hazard study 2 focussed on risk assessment and management over the plant as whole.
However, it is an on-going process and must be continually reassessed. We will regularly review the
hazards associated with our process and control them to an acceptable level. By taking these actions,
we ensure that we have done everything possible to prevent incidents, accidents and criminal actions.
Hazard study 3, also known as HAZOP, aimed to identify and evaluate deviations from the design intent
which would be hazardous and negatively affect operability. Unlike hazard study 2, which looked
across plant hazards in general, HAZOP focused specifically on individual units and all potential
dangers associated with them.
1.5. Company Policy, Legislation & Standards
The companyβs first priorities are the health & safety of people on site, whether they are employees
or visitors, and the welfare of the local environment. The policies upheld by Situs demonstrate our
duty as an employer to encourage an accident free working environment, along with minimising its
adverse effects on the local environment. The company policies are listed and discussed in greater
detail in Hazard Study 1, along with a list of key standards that Situs will adhere to and a list of
authorities to be contacted to ensure that these standards are met.
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2. Detailed Design: DR Flotation Cell
2.1. Introduction
The detailed design will focus on the use of the Denver DR Flotation Machine during the rougher
flotation stage of the process. The section will cover the theory and fundamentals of flotation and
apply the findings when sizing and designing the cell which will be a part of a bank of identical cells.
There are multiple established methods that can be used to size a flotation cell, some of which are
discussed in this section, however the chosen method of sizing involves modelling the cell as a mixing
tank in which all solid particles must be kept suspended in the air-brine mixture (Sinnott, 2005).
Flotation is the most important and versatile mineral separation technique, which usually takes place
in a water-mineral slurry, however in this process brine has replaced the water as the liquid phase.
The process is used to separate particles based on their surface chemistry. The surfaces of certain
minerals in the slurry are made hydrophobic by introducing conditioning reagents. These hydrophobic
particles become attached to air bubbles introduced to the slurry and rise to the surface, forming a
froth which can then be scraped off. There are a number of variables that affect the flotation recovery
rate as well as the addition of conditioning reagents. These include particle characteristics such as size
and shape, flotation machine variables such as equipment size and speed of operation, and finally
operating variables such as feed properties (King, 1982).
The most influential factor that affects the flotation recovery rate is retention time, it determines the
volume and number of flotation cells required. Therefore there are two alternative approaches when
sizing cells, having small cells and longer banks or fewer large cells and shorter banks. The latter is
considered more appropriate for high tonnage operations therefore this approach will be used as it
minimises capital and operating costs. Although the Reactor Cell System Flotation Machine is the
favoured choice for the majority of mineral flotation applications, the Denver DR Flotation Machine is
better suited to dealing with de-slimed coarse particles which are encountered in glass and potash
processing (Metso, 2008), a schematic diagram of the Denver DR Flotation Machine can be seen in
figure 2.1.
Figure 2.1: Schematic diagram of a Denver DR Flotation Machine (Sinnott, 2005).
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The slurry is introduced to the cell through a feed box and is distributed over the width of the first cell,
the remaining cells found on the same bank are separated using discharge boxes and intermediate
boxes in place to increase the hydraulic head to a level that will allow slurry to flow along the bank.
Each cell contains a near bottom impeller that subjects the slurry to intense agitation and aeration,
resulting in improved solid suspension and bubble dispersion (Sinnott, 2005). Each impeller is
surrounded by a diffuser to aid the dispersion of bubbles throughout the cell.
Low pressure air is introduced down the standpipe surrounding the impeller shaft and is dispersed
immediately in the form of small bubbles to ensure maximum contact with the KCl particles. Each unit
is contained in its cell by the removable froth baffles shown in figure 2.1 and creates an axial flow,
producing strong vertical flows at the base of the tank in order to suspend the heavier coarse particles
(Sinnott, 2005). The removable froth baffles placed between each cell prevent the transfer of froth
from one unit to another but still allow the liquid to flow freely through the bank, they also prevent
negative axial mixing effects by interrupting the flow of the liquid near the cell wall (Perucca, 2001).
2.2. Principles of Flotation
Flotation is a physical separation process involving three phases: solid, water, and air. However, as
previously stated the liquid phase being utilised in this process is brine. Designing and operating a
successful flotation system is a challenging task therefore gaining a clear understanding of the
fundamentals, such as the wettability of solid surfaces and the mechanical factors that affect recovery
rates, is essential.
2.2.1. Thermodynamics of Wetting
The primary objective of flotation is to bring the suspended solid particles into contact with air bubbles
to produce a stable bubble-particle attachment. The attachment of a solid particle to an air bubble
destroys the interfaces the brine has with the solid and air to create a new air-solid interface. The free
energy change, on a unit area basis is calculated using equation 2.2.1.
βπΊ = πΎ π΄π β (πΎππ΅ + πΎ π΄π΅)
(Perry, et al., 1997) (2.2.1)
Where πΎ represents the interfacial tensions of all three types of interface. A force balance for the air-
brine-solid particle system produces equation 2.2.2.
πΎ π΄π = πΎππ΅ + πΎ π΄π΅ cos π
(Perry, et al., 1997) (2.2.2)
Where π is the contact angle, shown in figure 2.2. The contact angle is an equilibrium measure of the
interfacial energy of the air-brine-solid system. Combining equations 2.2.1 and 2.2.2 produces 2.2.3.
βπΊ = πΎ π΄π΅(cos π β 1)
(Perry, et al., 1997) (2.2.3)
Therefore, for any finite value of the contact angle, the
free energy change will be negative meaning particle-
bubble attachment can take place. The wettability of solid
particle can be altered by the use of chemical reagents,
which are discussed in greater detail in section 2.3.
Figure 2.2: Contact angle between air bubble and
solid particle (Perry et al., 1997).
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2.2.2. Probability Process
It is possible to analyse the froth flotation recovery rate as a probability process, described by equation
2.2.4.
ππ = ππ ππ(1 β ππ)
(Ives, 1984) (2.2.4)
Where ππ , ππ , ππ , and ππ are probability of flotation , particle-bubble collision, particle βbubble
attachment and detachment, respectively. The combination of these three stochastic events produce
a final probability value which represents the likelihood of a particle being floated. The first term ππ is
controlled by the particle and bubble sizes, the number of bubbles and pulp density and the intensity
of agitation in the cell. The second term ππ depends on the value of the contact angle, discussed
previously and shown in figure 2.2. The third term ππ is dependent on the strength of the bond
between the particle and the bubble which can be easily broken due to liquid turbulence and impacts
from other suspended particles (Ives, 1984).
The relationship described above can be used as a method of sizing a bank of cells as the individual
probabilities can be calculated if particle and bubble sizes are known, however in this case accurate
estimates canβt be made, therefore the rougher flotation bank will be sized using a method found in
literature (Metso, 2008). Following this each individual cell will be modelled as a tank containing a
three phase mixture, in which all solid particles must be kept suspended. The method is discussed in
greater detail from section 2.5 onwards.
2.3. Reagents
There are three main types of reagents that are used during the flotation stage of the process:
collectors, frothers and modifiers (Zhang, 2012).
2.3.1. Collectors
Collectors are surface-active agents that are introduced to the slurry, and either bond chemically on a
hydrophobic mineral surface, or adsorb onto the surface. They increase the natural hydrophobicity of
the surface, aiding the efficient separation of the hydrophilic particles from the hydrophobic ones
(Leja, 1982). A classification of commonly used collectors can be seen in figure 2.3. Non-ionising
collectors are insoluble in water and cause the surfaces of particles to become hydrophobic by
covering them in a thin film. The ionising collectors disassociate into ions in water and are made up of
complex heteropolar molecules, they absorb either chemically or physically on the surface of the
particle and can be further categorised into anionic or cationic. Cationic collectors, especially the
primary and secondary amine salts, are used to float sylvite (KCl) from halite (NaCl) in brine solutions
(King, 1982). Dosage requirements for collectors depend on the mechanism by which they react with
the particle surface, but just enough is needed to form a monomolecular layer. The addition of excess
quantities of collector is
undesirable as it will result in
reduced selectivity and
increased material cost. The
collector being used in this
process is an alkyl-amine
called βFlotigam Sβ and will be
added at a rate of between 50
and 100 grams per tonne of
ore.Figure 2.3: Classification of collectors (Glembotski, et al., 1972).
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2.3.2. Frothers
Frothers are also surface-active agents and are added to the slurry in order to stabilize the air bubbles
and encourage particle-bubble attachment, carryover of particle-laden bubbles to the froth, and
removal of the froth (Leja, 1982). The frother action is primarily concentrated at the air-liquid interface.
The name of the frother being utilised in this process is βFlotanol C07β and is a polypropylene glycol, it
will be added at a rate of between 25 and 50 grams per tonne of ore (Kawatra, 2001).
2.3.3. Modifiers
Flotation modifiers include several classes of chemicals: activators, pH regulators, depressants,
dispersants and flocculants. Depressants aid the selectivity and prevention of unwanted minerals,
such as clay, from floating and in this process urea-formaldehyde is being used at a rate of between
30 and 60 grams per tonne of ore. Also being added during the conditioning stage is an extender oil
to facilitate the flotation of coarse particles (Perucca, 2001), along with a polyelectrolyte to block
adsorption sites on residual insoluble particles to prevent the adsorption of the potash collector
(Tippin, 1999).
2.4. Mass Balance
The mass balance across section four gives an overview of the total mass of each stream entering and
leaving each unit. The letters and numbers used refer to the process flow diagram in figure 2.4, a
description of the calculations and assumptions made throughout the mass balance are explained for
each unit following the process flow diagram. Streams 2, 4 and 11 do not appear in the mass balance
as stream 2 only contains the flotation reagents which are added in such a small quantity that they
can be ignored for these calculations, and streams 4 and 11 only contain the air introduced for the
flotation process which is assumed to exit to the atmosphere as the froth destabilises in the launder.
The units used for every stream is tonnes per hour as the basis of the mass balance is over a one hour
period. Appendix A shows the mass flow rate and mass fraction for each component of every stream.
Figure 2.4: Process Flow Diagram of section four.
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One assumption that is made throughout the process is that no material is lost. Therefore an overall
mass balance can be carried out to ensure that the total input equals the total output.
Total Input: 1687 tn/hr
Total Output: 1687 tn/hr
It is assumed that the size reduction section of the process reduced the maximum particle size down
to around 2500 microns, it also assumed that the particle sizes still follow a log normal distribution.
The units shown in figure 2.4 are labelled as follows:
4-A Conditioner 4-D Agitated Balance Tank 4-G Screen
4-B Rougher Flotation 4-E Cleaner Flotation 4-H Agitated Balance Tank
4-C Screen 4-F Sieve Bend Screen 4-I Hydrocyclone
2.4.1. Conditioner (4-A)
Entering the conditioner are two streams (streams 1 & 2), however as stream 2 only contains the
flotation reagents, which are added at a rate of around 200 grams per tonne of ore, it has been ignored
for the purpose of this mass balance. Therefore stream 3 leaves the conditioner with the same
composition as stream 1 when it entered, as can be seen in appendix A. It is also assumed that stream
1 enters from section three with the correct solid % for the rougher stage, which is 35 % solid by mass
(Metso, 2008).
2.4.2. Rougher Flotation (4-B)
An exact recovery rate for the rougher flotation stage has not been calculated, however it has been
assumed that 90% of the KCl that enters the bank is floated. It is often the case that NaCl particles and
insolubles can be floated accidentally, as they become entrapped by the KCl (Perucca, 2001). A small
percentage of the brine will also be removed in the froth phase, this value has been assumed to be
5%.
% of KCl leaving in stream 6 90%
% of NaCl leaving in stream 6 5%
% of insolubles leaving in stream 6 5%
% of brine leaving in stream 6 5%
2.4.3. Screen (4-C)
A particle size distribution for the KCl particles entering the screen can be seen in figure 2.5. As stated
previously, it is assumed that the maximum KCl particle size is roughly 2500 microns and that the
particles follow a log normal distribution. As can be seen on the graph, a screen size of 840 microns
would stop around 50% of the KCl particles. However a percentage of the NaCl particles would also be
stopped as they have a maximum size of around 2000 microns. A percentage of the insolubles and the
brine would also be stopped as they would be entrapped by the KCl particles. The MATLAB code used
to draw up the particle size distributions shown in this section, can be seen in appendix B.
% of KCl leaving in stream 7 50%
% of NaCl leaving in stream 7 10%
% of insolubles leaving in stream 7 5%
% of brine leaving in stream 7 5%
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2.4.4. Agitated Balance Tank (4-D)
In the first agitated balance tank enough brine is added through stream 9 to reduce the solid % by
mass of stream 10 down to 21%, as the % of solids in the feed to the cleaner flotation system greatly
affects its efficiency. The solid % of the feed into a cleaner flotation cell should be 60% of the value of
the solid % entering the rougher stage, which in this case was 35%.
2.4.5. Cleaner Flotation (4-E)
The same assumptions apply here that were used during the rougher flotation stage.
% of KCl leaving in stream 12 90%
% of NaCl leaving in stream 12 5%
% of insolubles leaving in stream 12 5%
% of brine leaving in stream 12 5%
2.4.6. Sieve Bend Screen (4-F)
Streams 13 and 5 are added together to create stream 14 which enters the sieve bend screen. As the
sieve bend screen has a separation of 1410 microns, only around 10% of the NaCl particles will be
stopped by the screen, along with a small percentage of insolubles and brine. The majority of the KCl
passing through 4-F are the larger KCl particles that did not float during the rougher stage, therefore
it is assumed that around 80% of them will be stopped by the 1410 microns screen.
% of KCl leaving in stream 15 80%
% of NaCl leaving in stream 15 10%
% of insolubles leaving in stream 15 5%
% of brine leaving in stream 15 5%
Figure 2.5: Cumulative frequency of KCl particle diameters in stream 6.
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2.4.7. Screen (4-G)
Screen 4-G has a separation of 500 microns and therefore stops around 65% of all of the NaCl particles
that enters it, as can be seen in figure 2.6. As the KCl particles are relatively large in comparison to the
NaCl particles it is assumed that around 80% of the KCl will be stopped by the screen. It is also once
again assumed that the screen will stop a small percentage of the insolubles and brine.
% of KCl leaving in stream 19 80%
% of NaCl leaving in stream 19 65%
% of insolubles leaving in stream 19 5%
% of brine leaving in stream 19 5%
2.4.8. Agitated Balance Tank (4-H)
In the second agitated balance tank steam 20 is the only stream entering and stream 21 is the only
stream leaving, therefore the only assumption that can be made is that all of the material that enters
the balance tank, exits the balance tank. The second balance tank is in place to ensure a constant feed
rate to the hydrocyclone, as mentioned in the process overview.
2.4.9. Hydrocyclone (4-I)
Assuming that the hydrocylone is almost 100% efficient and knowing that a cut off diameter of 55
microns is being used, 90% of the solids that enter the hydrocyclone leave via the underflow. However
it is assumed that not all of the brine will leave through the overflow, 5% will leave with the solids via
the underflow.
% of brine leaving in stream 22 95%
% of solids leaving in stream 23 90%
2.4.10. Converting Flow Rates
Throughout the sizing process, there are calculations which require the flow rate to be specified as a
volume per unit time rather than a mass per unit time. For these calculations it is assumed that the
brine used in the process is fully saturated and has a composition of 64% water, 25% NaCl & 11% KCl
with respective densities of 1000 kg.m-3
, 2170 kg.m-3
, and 1980 kg.m-3
.
Figure 2.6: Cumulative frequency of NaCl particle diameters in stream 18.
19. Page | 14
To convert the mass flow rates to volume flow rates insert the mass and density values of each
component into equation 2.4.1 individually:
ππ =
π π
π π
β
(2.4.1)
Where ππ is the volume of any given component, π π is the mass and π π is the density of that
respective component.
2.5. Rougher Flotation Basic Specification
As the flow rates through the rougher flotation bank have been calculated using estimations of the
recovery rate, the size and number of cells can be calculated using a basic sizing method. The selection
of the size and number of cells for each stage of the flotation circuit is made by a three step calculation.
The first step is the determination of the total flotation cell volume and can be calculated using
equation 2.5.1:
ππ =
π Γ ππ Γ π
60 Γ πΆ π
(Metso, 2008) (2.5.1)
Where ππ is the total flotation volume required in m3
, π is the feed flow rate in m3
.hr-1
, ππ is the
flotation retention time in minutes which usually is between four and six minutes for potash, in this
case six minutes has been used to ensure maximum efficiency. π is the scale up factor and in this case
is equal to one and πΆ π is the aeration factor included to account for air in the pulp and always equals
0.85 unless stated otherwise (Metso, 2008). The flow rate into the rougher flotation stage is 816 tn.hr-
1
, which can be turned into metres cubed per hour using the method described in section 2.4. A flow
rate of 816 tn.hr-1
equals 555 m3
.hr-1
. Therefore a total flotation volume of 65.3 m3
is required for the
rougher stage.
The second step is selecting the number of cells per bank. The typical number of cells found in a bank
for common mineral flotation duties can be found in literature, for potash the typical number is
between four and six, and in this case it was decided that five would be used (Metso, 2008). The
volume per cell could then be calculated using equation 2.5.2:
ππππ’ππ πππ πΆπππ =
πππ‘ππ πΉπππ‘ππ‘πππ ππππ’ππ π πππ’ππππ
ππ’ππππ ππ πΆππππ
(Metso, 2008) (2.5.2)
Dividing the total flotation volume required by the number of cells selected gives the volume per cell
to be 13.1 m3
, meaning the DR500 model will have to be used as it is the only one with a cell volume
greater than that. The DR500 model has a cell volume of 14.16 m3
and a diameter of 2690 mm.
The third step is selecting the bank arrangement. Intermediate boxes may be required to increase the
hydraulic head to a level that will allow slurry to flow along the bank. The maximum allowable numbers
of cells in a section between discharge is intermediate boxes can be found in literature, for the DR500
model the maximum number of cells per bank section is four, meaning at least one intermediate box
will be needed to split the five individual cells (Metso, 2008). The bank arrangement selected in this
case is F-3-I-2-D, figure 2.7 shows how the bank will be arranged from a side view and figure 2.8 shows
20. Page | 15
how the bank will be arranged from an aerial view. The detailed design of the DR flotation machine
will focus on the first cell in the bank arrangement, highlighted in figures 2.7 and 2.8.
The major dimensions for the rougher flotation system can be seen in table 2.1. The dimensions listed
in the table correspond to the dimensions shown in figures 2.7 and 2.8. The data was collected from
the Metso brochure for DR flotation cells (Metso, 2014). The value for D varies depending on by how
much the hydrostatic pressure of the fluid needs be increased upon entry to the next bank.
Table 2.1: Dimensions for rougher flotation system (Metso, 2014).
Model
Cell Volume A B C D (min) L W H
m3
mm mm mm mm mm mm mm
DR 500 14.16 762 914 914 305 2692 2692 3404
The section highlighted in blue in figure 2.9 represents the main walls of the cell, the specification for
which is given in section 2.6. The area highlighted in green is the diffuser and draft tube, specified in
sections 2.8 and 2.9 respectively. Their purpose is to encourage slurry recirculation by drawing large
volumes of material from the upper zone to break up any large concentrations of solids at the bottom
of the tank. It also encourages the bubbles to diffuse over the base of the cell and collide with solid
particles, rather than rising straight up to the top.
Figure 2.7: Side view of rougher flotation bank arrangement and dimensions for DR500 model (Metso, 2014).
Figure 2.8: Aerial Side view of rougher flotation bank arrangement and dimensions for DR500 model (Metso, 2014).
21. Page | 16
As the slurry will flow through multiple cells in one bank, removable baffles have to be fitted to prevent
froth migration between cells. The specification for these baffles, highlighted in red in figure 2.9, can
be seen in section 2.11. Another notable feature of the flotation cell displayed in figure 2.9 is the
launder, which is shown leading off the front and rear faces of the cell. Its purpose is to allow the
accumulating froth to drain freely off the top of the cell, as this design does not include a skimming
mechanism to remove it. Once the froth runs off the top of the cell it is collected in an underflow
system which is described in more detail in section 2.12.
2.6. Cell Wall Thickness
The cell wall thickness can be calculated using equation 2.6.1, used for calculating the thickness of a
flat plate required to resist a given pressure load. This equation is being used as the DR flotation cell
has a square base and is not cylindrical:
π‘ = πΆπ·β
π π·
π
(Sinnott, 2005) (2.6.1)
Where π‘ is the thickness, π is the design stress for the material being used, π· is the effective plate
diameter which in this case is the cell diameter, πΆ is a constant that depends on the edge support and
π π· is the design pressure. Assuming that SS-304 or SS-316 will be used to construct the walls of the
cell the design stress will be 220 MPa (BSSA, 2015). The design pressure can be specified by calculating
the pressure acting on the tank wall and increasing the value by 10% (Sinnott, 2005), as shown below:
π π· = 1.1(ππβ)
(Sinnott, 2005) (2.6.2)
Where π is the slurry density, π is acceleration due to gravity and β is the fluid height. The values used
to calculate the tank thickness are shown in table 2.2. The calculated thickness of each vessel wall was
15 mm, however a 2 mm corrosion allowance has been added, increasing the thickness to 17 mm.
Table 2.2: Values used in calculation of cell wall thickness.
Slurry
Density
Fluid
Height
Design
Pressure
Constant
Tank
Diameter
Design
Stress
Tank
Thickness
Ο (kg.m-3
) h (m) (Pa) C D (m) f (MPa) t (mm)
1643 2 35,460 0.43 2.7 220 17
Figure 2.9: Colour coded 3D diagram of a DR flotation cell (Yara, 2000).
22. Page | 17
2.7. Impeller & Air Delivery
The primary physical requirement of a flotation cell, regardless of what type it is, is to ensure the
dispersion of finely divided air bubbles throughout the cell. To do this effectively a flotation cell must
meet two main requirements; suspension and aeration (Yan & Gupta, 2006). The two key variables
that greatly affect suspension and aeration are impeller speed and air flow rate. This section will aim
to calculate the optimum operating conditions for the rougher floatation stage.
2.7.1. Suspension
It is essential that the impeller being used is capable of maintaining solid suspension during operation.
An insufficient level of agitation may cause the solids, especially the larger ones, to sediment (Yan &
Gupta, 2006). A degree of settling out will always occur in the corners of cells, however a significant
sedimentation of solid particles on the base of the cell can disrupt flow patterns and restrict the
required contact between air bubbles and the suspended particles. Particles that are not suspended
cannot make successful collisions with air bubbles. The DR principle of vertical recirculation of slurry
induced by axial mixing effectively combats stratification and sedimentation (Sinnott, 2005).
2.7.2. Aeration
The DR flotation machine has a controlled aeration rate, allowing it to be maximised independently of
impeller speed (Yan & Gupta, 2006). In some cases the rotating impeller causes the pressure near the
impeller to drop below the hydrostatic pressure in the pulp so that air may be sucked into the impeller
region. This is known as induced air and the practice of introducing supercharged air into the impeller
region is known as sub-aeration (Yan & Gupta, 2006). For this flotation process, only sub-aeration
takes place. It is important that the bubbles introduced are finely disseminated, and that there is
enough air to produce a stable froth of reasonable depth.
Cell performance is strongly related to the size of the particles being floated, and in this process the
range of sizes entering the rougher flotation stage is particularly large. The large range in the feed size
results in the optimum conditions for the flotation of the coarse particles to be considerably different
to the optimum conditions for the flotation of the fine particles (Yan & Gupta, 2006). Regardless of
the particle size, the speed of the impeller and the diameter of the air bubbles have the following
effects:
ο§ A low impeller speed will result in the particles not being suspended, but settling in significant
quantities at the base of the cell
ο§ A high impeller speed will result in turbulence that could be great enough to destroy the
bonds between solid particles and air bubbles
ο§ A low bubble size may result in bubbles not being able to offer sufficient buoyancy to float
solid particles and lift them through the slurry
ο§ A high bubble size will result in fewer bubbles being formed for a constant air flow rate,
resulting in a decreased recovery rate
2.7.3. Impeller Design
The design of the impeller and dip pipe shown in figure 2.10 creates the required axial flow needed
for vertical recirculation, therefore when choosing the impeller the type of flow it creates does not
need to be considered. In this case the popular 8 bladed Rushton Disk Turbine (RDT) is being used. The
ratio of impeller diameter to tank diameter is often between 0.3 and 0.5, in this case a value of 0.4
has been chosen; as the tank diameter is 2700 mm the chosen ratio gives an impeller diameter of 1080
mm.
23. Page | 18
The impeller can be situated as close as half the propeller diameter to the tank bottom without causing
a significant increase in the power drawn. The ratio of blade diameter to impeller diameter is 2:7 and
the ratio of blade width to impeller diameter is 1:7 (Lima, et al., 2009). This gives a blade diameter of
300 mm and a blade width of 170 mm. The width of the shaft connecting the impeller to the motor is
300 mm in diameter and the width of the pipe introducing the air into the centre of the impeller is
660 mm. Also connected to the impeller is the diffuser and draft tube, discussed in section 2.5. The
dimensions of these two parts of the unit are specified in sections 2.8 and 2.9.
The Zweitering correlation was constructed to calculate the impeller speed at which off bottom
motion began to occur in a solid suspension in an agitated vessel (Zwietering, 1958). At this impeller
speed the maximum surface area of the solid particles is exposed for mass transfer. The original
correlation only applied to solid-liquid systems. However, recently Van der Westhuizen and Deglon
investigated the critical impeller speed in a mechanical flotation cell and adapted the correlation so
that the air flowing through the system could also be accounted for. Equation 2.7.1 describes how the
critical impeller speed ππ½π, can be calculated for a mechanical flotation cell:
ππ½π = πΎππΏ π π
0.33
π0.17
(
π π β π πΏ
π πΏ
)
0.70
(
π πΏ
π π
)
0.05
(1 + π πΊ π πΊ)
(Lima, et al., 2009) (2.7.1)
Where: ππ½π is the critical impeller speed (s-1
)
πΎππΏ is a constant related to impeller size and design
π π is the particle diameter (Β΅m)
π is the mass fraction of solids in suspension
π π is the average density of the solid particles (kg.m-3
)
π πΏ is the density of the liquid (kg.m-3
)
π πΏ is the kinematic viscosity of the liquid (m2
.s-1
)
π π is the kinematic viscosity of water (m2
.s-1
)
π πΊ is a constant related to the impeller (s.cm-1
)
π πΊ is the superficial gas velocity (m.s-1
)
Figure 2.10: Significant dimensions of impeller dip pipe (Lima, et al., 2009).
24. Page | 19
The constant πΎππΏ can be taken from literature and has a value of 22 for this type and size of impeller.
For a log normal distribution with a maximum particle size of 2.5 mm, the most frequent particle size
is 350 microns, therefore this will be used as the value for π π. The value was calculated using the
MATLAB code displayed in appendix B. The mass fraction of solids in suspension (π) can be calculated
using the data from the mass balance in appendix A as it is equal to the mass fraction of solids in
stream 3, which is 0.35. The average density (π π) of the NaCl and KCl is 2096 kg.m-3
and the density of
the fully saturated brine solution (π πΏ) is 1400 kg.m-3
.
The kinematic viscosity π πΏ is calculated by dividing the brines dynamic viscosity (0.005 Pa.s for this
fully saturated brine solution) by the density, which is 1400 kg.m-3
. This relationship is represented by
equation 2.7.2:
π πΏ =
π πΏ
π πΏ
β
(Perry, et al., 1997) (2.7.2)
Where π πΏ is the dynamic viscosity of the liquid. This gives a value of for the kinematic viscosity of the
liquid as 3.57 x 10-6
m2
.s-1
. The kinematic viscosity of water π π is 6.39 x 10-7
m2
.s-1
and the constant
π πΊ is equal to 0.40 s.cm-1
for this impeller type. The superficial gas velocity π πΊ is calculated by dividing
the gas flow rate into the cell by the cross sectional area of it, as shown in equation 2.7.3:
π πΊ =
π πΊ
π΄
(Perry, et al., 1997) (2.7.3)
Where π πΊ is the gas flow rate in m3
.s-1
and π΄ is the cross sectional area of the tank in m2
. The air
requirement of a DR500 flotation cell is 6.5 m3
.min-1
which equates to 0.11 m3
.s-1
. Dividing 0.11 m3
.s-1
by the tank cross sectional area, 7.29 m2
,gives a superficial gas velocity of 0.0151 m.s-1
.
Inserting all of these values into equation 2.7.1 gives a critical impeller speed of 85 RPM. However,
the best operating conditions are 10% above ππ½π (Simmons, 2013), and for that reason the impeller
will run at around 94 RPM.
2.7.4. Air Requirements
As previously mentioned the air requirement of a DR500 flotation cell is 6.5 m3
.min-1
supplied at a
gauge pressure of 18 kPa (Metso, 2008). Each cell in the bank is provided with an individually
controlled air valve (Sinnott, 2005). The type of valve controlling the quantity of air delivered into each
unit is a butterfly valve, which is discussed in greater detail in section 3.9. The air is delivered into the
system through a dip pipe surrounding the impeller shaft, which stretches vertically all the way up
through the slurry and froth and is connected to the plants main air supply. The dip pipe is 660 mm in
diameter, which can be seen in figure 2.10. As shown in figure 2.1 the pulp and air meet and mix in
the open throat of the impeller, before being ejected by the impeller over the base of the cell.
2.7.5. Power Requirements
Now that the operating impeller speed, air requirements and all of the major tank and impeller
dimensions have been specified the power requirements of the cell can be calculated. Before
calculating the power drawn, the type of flow in the cell needs to be determined i.e. if the flow is
laminar or turbulent. The Reynolds number can be calculated using equation 2.7.4. If the Reynolds
number is greater than 20,000 then the flow in the vessel is turbulent:
π π =
π πΏ ππ·2
π πΏ
(Simmons, 2013) (2.7.4)
25. Page | 20
Where π πΏis the density of the liquid which is 1400 kg.m-3
, π is the impeller speed which is 1.57 rev.s-
1
, π· is impeller diameter which is 1.08 m and π πΏ is the dynamic viscosity of the liquid which is 0.005
Pa.s. Inserting these values into equation 2.7.4 produces a Reynolds number of around 5 x 105
which
is well into the turbulent range. Following this, equation 2.7.5 can be used to calculate the power
drawn by the impeller:
ππΌ = ππ(π πΏ π3
π·5)
(Simmons, 2013) (2.7.5)
Where ππ is the power number which for a Rushton disk turbine operating in turbulent flow is 5.
Therefore the calculated power drawn by the impeller is 39.6 kW.
2.7.6. Motor Design
The power required will be delivered by an electric motor attached to the top of the impeller shaft.
Electric motors tend to operate at standard speeds of 1800 or 1500 RPM, which is a lot faster than
necessary for this process (WEG, 2014). Therefore, gearboxes will also be fitted to reduce the number
of revolutions per minute but increase the torque provided.
The type of electric motor that will be used is called an induction motor. This choice has been made
as they generally run at a constant speed except for when sudden large mechanical loads are applied
to the motor shaft. It also has a simple design, is robust, cheap and suitable for almost all types of
machinery. Additionally the speed at which the induction motor runs can be controlled by the use of
frequency inverters (WEG, 2014).
2.8. Diffuser
The role the diffuser plays during flotation is discussed in detail in section 2.5, and the position it takes
up inside the cell can be seen in figure 2.9. Typical ratios of diffuser diameter to impeller diameter
were discovered in literature (Lima, et al., 2009) and it was found that the width of the diffuser is only
marginally larger than the impeller diameter, but smaller than the draft tube diameter.
For these reasons a diameter of 1200 mm will be used for the diffuser. The diffuser will attach to the
bottom of the dip pipe as shown in the mechanical drawings in appendix C.
2.9. Draft Tube
The role the draft tube plays during flotation is discussed in detail in section 2.5, and the position it
takes up inside the cell can be seen in figure 2.9. Typical ratios of the draft tube compared to the
diameter of the impeller have been found in literature (Lima, et al., 2009) and it was discovered that
the upper width of the draft tube roughly equalled the width of the diffuser and the lower width of
the draft tube was only marginally larger than the diffuser diameter.
For these reasons a diameter of 1200 mm will be used for the upper width of the draft tube and a
diameter of 1400 mm will be used for the lower width of the draft tube. The dimensions of the draft
tube and the way the draft tube will be attached to the dip pipe is shown in the mechanical drawings
of the unit in appendix C.
26. Page | 21
2.10. Inlet
An inlet box is always positioned at the start of a bank of flotation cells. As the flow into the bank may
fluctuate the inlet box can hold a percentage of the material and steady the flow through the bank.
Pumping the slurry straight into a flotation cell may also disrupt the desired axial flow created by the
impeller, an inlet box also counteracts this problem.
The inlet box has a width of 2692 mm, which is the same as the width of the cell. The point at which
the pipe enters the inlet box must be greater than the height of the slurry to avoid having to pump
the slurry into the flotation bank. Therefore the bottom of the entry point has a height of 1600 mm.
The inlet box has a length of 762 mm as specified in section 2.5. All of the dimensions calculated can
be seen in figure 2.11.
2.11. Removable Baffles
The thickness of these removable baffles is only a
third of the cell wall thickness, 5 mm to be exact, as
the pressure exerted on them by the froth is not as
great as the pressure exerted by the slurry, due to
the large different in densities. They have the same
width as the cell, 2692 mm, as they have to stretch
all of the way across to prevent any froth migration.
Their depth will also have to be greater than that of
the froth. A typical height of froth in a flotation
vessel is 15% of the height of the slurry (Wills &
Napier-Munn, 1985). The height of the slurry in the
vessel can be calculated by dividing the volume of
slurry in the cell by the cross sectional area of the
cell.
The volume of slurry in each cell can be calculated by
first finding the mass of slurry that passes through
each cell every six minutes, the length of the residence time, which can be found in the results from
the mass balance in appendix A. The mass value can then be converted into a volume using the
conversion method described at the end of section 2.4. The volume of slurry in the first cell of the
Figure 2.12: Major dimensions of removable baffles.
Figure 2.11: Major dimensions of inlet box.
27. Page | 22
bank at any given time was calculated to be 9.57 m3
, which equates to a slurry height of 1.31 m.
Multiplying this value by 0.15 would give the height froth to be around 0.2 m.
The baffles will sit 500 mm above the floor of the cell, therefore a baffle depth of 1200 mm would be
appropriate as this would also allow for a slight fluctuation in the slurry due to varying flow rates or
froth height. The calculated dimensions of the removable baffles can be seen in figure 2.12 above.
2.12. Launder & Trough
Launders are fitted to a flotation cell to allow the froth to be removed. In some flotation cells a
skimmer is used to remove the froth from the top, however in this design the froth is being allowed
to flow freely over the sides of the vessel before being caught by the troughs that lie on either side as
shown in figure 2.13.
The launders will have a width of 320 mm and stretch the entire length of the cell and the trough will
have a width of 540 mm. The trough is sloped so that the extracted froth will drain freely before being
sent to the centrifuge for drying.
2.13. Supports
2.13.1. Base Support
As the bank of cells all have flat square faced bases,
no supports are needed to raise them above the
ground. Instead they will rest on a raised concrete
block that has the same dimensions as the base of the
bank. The concrete block being used as the support
will have a depth of 500 mm. It is necessary to raise
the cell off of the ground slightly as the froth
extraction trough will have to run parallel to the cell
walls for the length of the bank. Figure 2.14 shows a
basic diagram of the inlet box and the first cell in the
rougher flotation system (without the impeller & air
delivery unit), along with the dimensions of the
concrete block the bank will rest on.
Figure 2.13: Major dimensions of launder and froth removal system.
Figure 2.14: Major dimensions of supports for flotation bank.
28. Page | 23
2.13.2. Shaft & Motor Mount
The impeller, shaft and motor unit will be mounted on a steel bar that runs parallel to the cell walls
and rest on top of the unit as shown in figure 2.14. It will have a width of 300 mm and a depth of 300
mm as this size bar will easily support the weight of the impeller unit.
2.14. Materials of Construction
The slurry that is being processed throughout section four contains large amounts of NaCl. When NaCl
dissolves in solution, chloride ions are produced. Chloride ions can cause high levels of corrosion in
the presence of steel, which is the most frequently used engineering material. For this reason, the
majority of the equipment across section four, including the DR flotation machines, will be constructed
from stainless steel.
Stainless steels are the most commonly used corrosion resistant materials in the chemical industry.
To have corrosion resistant properties the content of chromium in the stainless steel must be greater
than 12%, and the higher the chromium content is the more resistance to corrosion the material offers
in oxidising conditions. Nickel is also added to increase the resistance to corrosion in non-oxidising
conditions (Sinnott, 2005).
2.14.1. Forms of Corrosion
The most common forms of corrosion in stainless steel at ambient temperatures and pressures are:
Pitting corrosion describes localised corrosion that results in the formation of pits in metal surfaces.
Pitting corrosion occurs when the surface finish on a material is poor or when the material used
contains impurities (Sinnott, 2005).
Erosion-corrosion occurs in fluid streams that contain suspended particles or where there is fluid
moving at a high velocity. The risk of erosion-corrosion is high near the impeller as there is a large
number of suspended coarse particles moving around with very high velocities (Sinnott, 2005).
Galvanic corrosion can occur when dissimilar metals are placed next to each other in the presence of
an electrolyte. There is no risk of galvanic corrosion between different types of stainless steels,
however there is a slight risk of galvanic corrosion if stainless steel is placed next to mild steel (BSSA,
2015).
2.14.2. Materials
The materials being used to construct the flotation machine are:
SS-304 is the most widely used stainless steel. It contains the minimum amount of chromium and
nickel that gives a stable austenitic structure (Sinnott, 2005). It offers some level of resistance to
chloride ion containing solutions, but not enough to protect the material completely. For this reason
the only part of the unit constructed from SS-304 is the impeller mount, as it is the only part of the
unit that does not come into direct contact with the chloride ion rich slurry.
SS-316 has a similar composition to SS-304, except for the addition of molybdenum which is added to
improve corrosion resistance in reducing conditions, such as in solutions containing chlorides (Sinnott,
2005). The majority of the unit will be constructed from SS-316, as shown in table 2.3.
Rubber offers good resistance erosion-corrosion, and for that reason it will be used to construct the
areas of the unit that are most susceptible to it, as shown in table 2.3. As these parts of the unit are
being lined with rubber there is no need for the primary construction material to be stainless steel, as
its surface will not be exposed to the chloride ions. Mild steel will be used instead as it is much cheaper
29. Page | 24
than stainless steel. Due to the rubber lining that covers the mild steel, galvanic corrosion between
the mild and stainless steel will not occur (BSSA, 2015).
Table 2.3: Primary and secondary materials used to construct DR flotation cell.
Unit Part Primary Material Secondary Material
Cell Wall SS-316 -
Inlet SS-316 -
Baffles SS-316 -
Impeller Shaft SS-316 -
Dip Pipe SS-316 -
Impeller Mild Steel Rubber
Diffuser Mild Steel Rubber
Draft Tube Mild Steel Rubber
Impeller Mount SS-304 -
Launder SS-316 -
Froth Extraction SS-316 -
2.14.3. Design for Corrosion Resistance
As well as material section, the design of a plant can affect how resistant it is towards corrosion. The
life of equipment that may be exposed to corrosive environments can be greatly increased by proper
attention to design detail. Designing equipment so that it drains completely and freely will reduce its
exposure to corrosive materials. Ensuring the internal surfaces of units and pipes are smooth and free
from crevasses will reduce the chance of corrosive materials accumulating on them. Fluid velocities
should also be high enough to avoid the deposition of solids on material surfaces, but not so high as
to cause erosion-corrosion (Sinnott, 2005).
Although the detailed design is only being carried out for one of the DR flotation cell, the materials
selected in this section apply to all of the cells found in the rougher flotation stage of the process.
2.15. Mechanical Drawings
Mechanical drawings showing the top and side view of the DR Flotation Machine designed in this
section can be seen in appendix C. The mechanical drawings show all of the major dimensions
calculated in this section.
30. Page | 25
3. Additional Specifications
Section three gives basic specifications such as dimensions and the materials of construction for all
other units, valves, pumps, pipes and conveyors found in section four. It also describes the methods
used to size the equipment along with reasoning behind various selections. Completed data sheets
and basic diagrams for each type of unit can be found in Appendix D.
3.1. Cleaner Flotation
The cleaner flotation cells were sized using the same three step method and equations explained in
section 2.4 except for a few alterations. The flow rate π into the cleaner flotation stage is 302 tn.hr-1
which equates to 225 m3
.hr-1
. For cleaning applications the retention time ππ is reduced to 65% of the
original time and the amount of solids in the feed is also reduced from 35% by mass to 21% as
mentioned in the mass balance in section 2.3 (Metso, 2008). The scale up factor π remains as one and
the aeration factor πΆ π remains as 0.85. Inserting these numbers into equation 2.4.1 calculated a total
flotation volume of 17.2 m3
.
The number of cells per bank also remained the same, staying at five. Dividing the total flotation
volume by this value produced a volume per cell of 3.4 m3
, meaning the DR180 model will have to be
used as it is the next available size with a maximum bank feed rate greater than the flow rate of the
feed. A completed data sheet containing all of the dimensions and important details about the DR180
model along with a schematic diagram can be viewed in appendix D. The maximum number of cells
per bank section for the DR180 model is six, meaning the five required can be placed consecutively
without being separated by an intermediate box. Therefore the bank arrangement will be F-5-D. The
air requirement for a DR180 floatation machine is 3.1 m3
.min-1
which equates to a gauge pressure of
14 kPa.
The cleaner flotation bank will be constructed from the same materials as the rougher flotation bank.
3.2. Screens
Screening is the separation of a mixture of various sizes of particles into two or more portions by
means of a screening surface. Any particle that remains on a given screening surface is called the
oversize, whereas the material that passes through is called the undersize (Perry, et al., 1997). The
types of screening available and the range of separations that can be achieved with various screens
are shown below in figure 3.1 (Matthews, 1972). As all three screens in section four have a separation
that is reasonably close to 1mm, vibrating single inclination screens are suitable to use for all three
units. However, as general industrial practice dictates that potash rougher tails should be separated
using a sieve bend screen (Perucca, 2001), an exception will be made.
Figure 3.1: Range of separations that can be obtained with various screen types (Matthews, 1972).
31. Page | 26
When the material is passed through a single inclination screen at an angle of 15Β° the particles undergo
a circular motion leading to screening by stratification, this allows the fine particles to pass between
the larger ones resulting in a sharp separation (Metso, 2008). The required area of a screen can be
estimated based on the through flow rate of solids using equation 3.3.1.
π΄ =
0.4πΆπ‘
πΆ π’ πΉππ πΉπ
(Matthews, 1972) (3.3.1)
Where π΄ is the area of the screen, πΆπ‘ is the flow rate of solid material through the screen, πΆ π’ is the
unit capacity, πΉππ is the open-area factor and πΉπ is the slotted-area factor. The unit capacity of a screen,
πΆ π’, depends on two variables; the material being screened and the size of separation. In this case the
values were determined using experimental data (Matthews, 1972). The open-area factor for a screen
with standard square openings is calculated using equation 3.3.2. The slotted-area factor for a screen
with standard square openings is equal to one for all three screens, as it is effectively the length to
width ratio of the openings.
πΉππ = 100 (
π
π + π
)
2
(Matthews, 1972) (3.3.2)
Where π is the clear opening diameter of the holes in the deck and π is the wire diameter. A
recommended nominal wire diameter, π, can be found for each particular mesh size in literature
(Perry et al., 1997). Note that the value of π is 15% larger than the required separation as the screens
are constructed from polyurethane and on an incline of 15Β°.
The calculated screen areas can be found below in table 3.1. The largest screen size available is 14.4
m2
, meaning multiple screens will be required to run in parallel to process the total through flow rate,
therefore also shown in the table is the number of screens required for all three operations.
Completed data sheets for all three screen variations can be found in Appendix D, they include major
dimensions and the power consumed by the units during operation.
Table 3.1: Calculated screen areas and the number of screens required.
Parameters Units 4-C 4-F 4-G
Separation Β΅m 840 1410 500
Screen Type - Single Inclination Sieve Bend Single Inclination
Mesh Size - 20 14 32
Screen Area (A) m2
21.7 26.8 75.4
Through Flow Rate (Ct) tn.hr-1
130 212 350
Unit Capacity (Cu) tn.(hr.m2
)-1
0.056 0.062 0.047
Open-Area Factor (Foa) - 42.8 51.0 39.5
Slotted-Area Factor (Fs) - 1 1 1
Opening Diameter (a) Β΅m 966 1622 575
Wire Diameter (d) Β΅m 510 650 340
No. of Screens Required - 2 2 6
32. Page | 27
The screens are constructed from polyurethane as they favour wet screening for any size particle and
provide accurate screening. The polyurethane screens will also have a longer lifetime as they are less
susceptible to erosion-corrosion. The rest of the unit will be constructed from mild steel, however the
parts of the unit that come into contact with the chloride rich slurry will be coated with natural rubber.
3.3. Hydrocyclone
Hydrocyclones separate solids by mass using the
effect of centrifugal forces. They are extremely
popular in industry as they have a very low capital cost
and have the ability to make very fine separations. A
Hydrocyclone consists of a top cylindrical section and
a lower conical section that terminates in an apex
opening as can be seen in figure 3.2. The unit operates
under pressure induced by a pump on the inlet
stream (Perry et al., 1997).
Larger particles are removed in the underflow and
stay close to the outer wall of the unit, whereas the
smaller ones remain close to the centre before being
removed in the overflow. Although hydrocyclones
are typically used for size control they can also be
used for dewatering, thickening, desliming and
washing (Metso, 2008). In this stage of the process
the hydrocyclone is being used as a dewatering unit, allowing the solids removed to be sent to backfill
and the recovered brine to be recycled. To ensure efficient classification the amount of solids in the
feed must be kept to a minimum. A hydrocyclone can achieve good efficiency when the % solids by
volume is between 10% and 15% (Metso, 2008).
In order to size a hydrocyclone d50 must first be calculated. Most end users of cyclones donβt calculate
the value d50, in reality the selection is based on size analysis of the overflow. In this case it is known
that hydrocyclone is required to remove 90% of the solids that enter via the underflow, leaving 10%
of the feed solids to leave in the overflow. The cumulative frequency of particle diameters entering
the hydrocyclone is shown in figure 3.3, it
indicates that the smallest 10% of solids will
be 55 microns or smaller. Multiplying the
cut off diameter of 55 microns by an
efficiency factor can produce an accurate
estimate of the d50 value. For a separating
efficiency of 99% a factor of 0.49 is used
(Metso, 2008), this produces a d50 value of
27 microns.
Figure 3.4 shows that a d50 value of 27
microns can be achieved using a cyclone
with a diameter of 420 mm. The feed enters
the hydrocyclone at a rate of 1180
tonnes.hr-1
which equates to 910 m3
.hr-1
.
Figure 3.4 shows that a cyclone diameter of
420 mm can achieve a flow rate of around
Figure 3.2: Schematic diagram of a hydro cyclone
(Metso, 2008).
Figure 3.3: Size of particles entering hydrocyclone.
33. Page | 28
300 m3
.hr-1
, meaning that three hydrocyclones will need to be run in parallel to achieve the required
degree of separation. The hydrocyclones operating pressure will vary between 120 kPa and 150 kPa.
It will be constructed of SS-304 but lined with rubber (Weir Minerals, 2008).
3.4. Agitated Balance Tanks
Balance tanks are an important part of many processes as they can give a constant pump head or flow
rate through a sequential unit. As hydrocyclone efficiency is greatly dependant on the constant flow
of material through it a balance tank will be placed leading up to it, another balance tank will also be
positioned before the cleaner flotation stage. The balance tanks used in this process will also be
agitated to ensure good homogeneity when the material exits the vessel, as the solid particles may
sediment rapidly due to their weight and size. Equation 3.5.1 was used to calculate the required
volume of each balance tank, a retention time of an hour was used as this would give time for any
minor maintenance to be carried out on sections of the plant without having to shut down the entire
operation. It was also found that a residence time of an hour would leave a greater than sufficient
amount of time to homogenise the slurry.
ππππ’ππ π πππ’ππππ = πΉπππ€ π ππ‘π Γ π ππ‘πππ‘πππ ππππ
(Metso, 2008) (3.5.1)
Once the required volume had been calculated the closest available volume was selected from the
literature, however an increased level of 20% was designed for as the tank should never be completely
full (Metso, 2008). As the tank depth to diameter ratio was close to or equal to one a Single MIL
impeller was selected to agitate the mixture. The remaining details found in the literature can be
viewed in the completed data sheets in appendix E. The majority of the vessel will be constructed from
SS-316 and the mounts will be constructed from SS-304. However, the impeller will be constructed
from mild steel but lined with rubber to prevent erosion-corrosion from the coarse particles.
Figure 3.4: Acceptable d50 values and flow rates for given cyclone diameters (Metso, 2008).
34. Page | 29
Table 3.2: Significant values & dimensions of balance tanks.
Tank Parameters Units 4-D 4-H
Flow Rate
tn.hr-1
302 1208
m3
.min-1
3.8 15.4
Retention Time min 60 60
Volume Required m3
225.0 924.0
Tank
Diameter m 8 12
Height m 7 12
Volume m3
317 1221
Impeller
Type - Single MIL Single MIL
Diameter mm 3050 4570
No. of Blades - 6 6
Motor Power kW 30 75
3.5. Conditioner
Before material can be sent through the flotation stage it must first be conditioned, this is done by the
addition of flotation reagents, which were discussed in greater detail in section 2.3. The conditioner
used for this process was sized using the same equation that was used to size the agitated balance
tanks, as the principle of mixing and retention time is very similar in both cases. The same impeller
used for the balance tanks will be used for the conditioner, however it will also be fitted with a draft
tube to prevent the material short circuiting the internal baffles. Due to the large size of the solid
particles it is assumed that a heavy duty mechanism will be required to ensure solid suspension. All
other major properties of the conditioner can be found in table 3.3, any data that was not calculated
was extracted from specifications given in the literature (Metso, 2008). A completed data sheet and
schematic diagram for the conditioner can be found in appendix D. The majority of the vessel will be
constructed from SS-316 and the mounts will be constructed from SS-304. However, the impeller will
be constructed from mild steel but lined with rubber to prevent erosion-corrosion from the coarse
particles.
Conditioner Parameters 4-A
Flow Rate
910 tn.hr-1
10.3 m3
.min-1
Retention Time 60 min
Volume Required 555 m3
Tank
Diameter 8 m
Height 7 m
Volume 317 m3
Impeller
Type Single MIL
Diameter 2745 mm
No. of Blades 6
Motor Power 30 kW
Table 3.3: Properties of conditioner
35. Page | 30
3.6. Conveyors
In areas of the plant where the solid % of the stream is too high for the material to be transported by
pipeline, an alternative method will be used. If the amount of solids in a stream is greater than 70 %
by mass then a conveying system will be used in place of a pipe (Metso, 2008). This cut off point results
in five separate conveyors being utilised across the plant, these streams can be seen in table 3.4 along
with their flow rates. Conveyors are selected from five key parameters; tonnage, material, size,
inclination and distance. Conveyor belts have a maximum capacity of 350 tn.hr-1
and a maximum feed
size of approximately 50mm, meaning they would be suited to deal with all five streams (Metso, 2008).
All five streams contain a significant percentage of moisture meaning that dust emission will not be a
problem, there is also no transportation of hot materials which eliminates further risks that wouldβve
had to be considered. Flat belts can carry materials up to distances of around 500m and convey up to
a lifting or lowering angle of 18Β°, which is also the
maximum inclination at which potash can be conveyed
(Metso, 2008). As none of the units that are attached to
the conveyors are exceptionally large in height or far
away from one another flat belts will be used for all five
streams. The frames supporting the belt are
constructed from mild steel, however the belts
themselves are made from rubber and reinforced with
polyester, polyamide, aramid and steel chords (Metso,
2008).
3.7. Pipelines
The pipe for which a design specification will be given has been chosen as stream 20, which runs
between unit 4-G and 4-H. A manual valve (MV-20/03) and centrifugal pump (CP-20/01) that are
placed on stream 20 will also be designed in the following sections. The method described will be used
by Situs to size all pipes used for the process.
All of the pipes in section four will be constructed from mild steel and lined with Rubber βaβ (Trellex
T40) to avoid the risk of erosion-corrosion (Metso, 2008). The wear rate of stainless steel pipes is
roughly 1.29 mm/year, however Rubber βaβ (Trellex T40) has a wear rate of only 0.13 mm/year, giving
it a life expectancy 10 times greater than that of stainless steel (Metso, 2008).
The pipes will also be constructed in a way that allows them to drain freely, to avoid any deposition of
materials for long periods of time. All of the pipes in section four will be Schedule 40 as it is the
standard size used in industry (Sinnott, 2005).
As stream 20 follows on from the underflow of a 0.5 mm screen, all of the particles in the slurry will
be below this size. Therefore the average particle diameter will be around 0.2 mm assuming the
particle sizes still follow a log normal distribution. For a slurry containing particles that have a diameter
between 0.1mm and 1mm an optimum pipe velocity of 2 m.s-1
should be used during calculations
(Abulnaga, 2002). The internal pipe diameter can be calculated using equation 3.7.1:
ππ = β
1.274 Γ π
π’
(Sinnott, 2005) (3.7.1)
Where ππ is the internal diameter of the pipe (m), π is the volumetric flow rate (m3
.s-1
) and π’ is the
optimum velocity for the fluid contained by the pipe (m.s-1
). The mass flow rate of stream 21 is around
Stream No.
Flow Rate
tn.hr-1
m3
.hr-1
6 150 84
7 58 27
8 92 53
12 62 42
19 285 152
Table 3.4: Flow rates of various streams.
36. Page | 31
1210 tn.hr-1
which equates to a volumetric flow rate of 0.198 m3
.s-1
when using the conversion method
described in section 2.4. Inserting this value into equation 3.7.1 along with the optimum velocity of 2
m.s-1
calculates the internal pipe diameter to be 355 mm.
Pipe size is specified with two non-dimensional numbers. The first is a nominal pipe size, abbreviated
to DN in Europe, for diameter. The second is the pipe schedule which is used for wall thickness, as
previously mentioned this pipe is schedule 40. The closest standard pipe size to 355 mm is DN 350
(Saylor, 2014). For DN 350 pipe with a schedule of 40, the outer diameter is 355.6 mm and the pipe
thickness is 9.525 mm, however 3 mm will be added to allow for the rubber lining, giving a total
thickness of roughly 12.5 mm. This gives an internal pipe diameter of 343 mm, which is slightly lower
than the calculated internal diameter. Therefore the velocity in the pipe will be slightly larger than 2
m.s-1
.
It is assumed that the length of pipe running between unit 4-G and 4-H is 8 m and has a total elevation
of 5 m. Situs understand that at this stage this is a conservative estimate and that values for pipe
length could change.
3.8. Pumps
The pumps in section four will all be centrifugal pumps and will all be positioned as close to upstream
equipment as possible to avoid cavitation. Centrifugal pumps have been chosen as they are well suited
to dealing with fluids containing a high percentage of solid material (Sinnott, 2005). The method used
for specifying pumps can be seen below. The pump specified in this section is CP-20/01. Potash
processing plants usually contain two pumps operating in parallel, both of which are able to pump the
full load (Perucca, 2001). Parallel pumps have been utilised across section four.
The nominal pipe diameter of the pipe that the pump is connected to is 350 mm, which gives a fluid
velocity of 2.04 m.s-1
for the desired flow rate. The first stage of the pump specification involves
calculating the losses due to friction. This can be done using equation 3.8.1:
βππ = 8π (
πΏ π
ππ
)
ππ π’2
2
(Sinnott, 2005) (3.8.1)
The friction factor π was calculated using the Moody chart. To use the Moody chart, the Reynolds
number and the relative roughness needed to be calculated, this was done using equations 3.8.2 and
3.8.3 respectively, the nomenclature and values for which can be found in table 3.5:
π π =
ππ π’ππ
π π
(Sinnott, 2005) (3.8.2)
π = π
ππ
β
(Telford, 2006) (3.8.3)
The absolute roughness (π) of a rubber lined pipe was taken to be 0.00015 (Abulnaga, 2002). The
viscosity of a slurry π π can be calculated using the following equations, the nomenclature and values
for which can be found in table 3.5:
π π = π π π πΏ
(Abulnaga, 2002) (3.8.4)
π π = 1 + 2.5β + 10.05β 2
+ 0.00273π16.6β
(Abulnaga, 2002) (3.8.5)
37. Page | 32
Table 3.5: Calculation of pressure drop in stream 20 due to frictional losses.
Symbol Parameter Units Value
ππ Slurry Density kg.m-3
1480
π’ Fluid Velocity m.s-1
2.04
π π Reynolds Number - 165,800
ππ Inside Pipe Diameter mm 350
π π Slurry Viscosity Pa.s 0.0063
π πΏ Liquid Viscosity Pa.s 0.0050
π π Relative Viscosity Pa.s 1.27
β Volume Fraction - 0.08
π Absolute Roughness m 0.00015
π Relative Roughness - 0.00043
π Friction Factor - 0.018
πΏ π Pipe Length m 8
βπ· π Pressure Drop to Friction Pa 10,100
There will be an additional pressure drop due to losses in bends and valves. The flow in the pipe is
turbulent, therefore the total loss can be approximated using tables found in the literature (Perry et
al., 1997). Between the pump and the downstream unit it is assumed there will be two 90Β° bends and
one manual knife gate valve. The losses can be measured in terms of velocity heads. The loss in velocity
heads through a ΒΌ open gate valve is 16 and for a 90Β° bend the loss 0.8, giving a total loss of 17.6
velocity heads (ππ£β) (Sinnott, 2005). This value can be converted into a pressure using equation 3.8.6:
βππ΅ = ππ π (
π’2
2π
ππ£β)
(Sinnott, 2005) (3.8.6)
Inserting the known values into equation 3.8.6 calculates the pressure drop due to fittings (βππ΅) to be
54,000 Pa. The difference in operating pressure between the two units (βπ ) is zero as they are both
operating at atmospheric pressure. The values calculated above can then be inserted into equation
3.8.7 to calculate the work done on the material per kilogram. If the value is negative, then a pump is
required.
π =
βπ
ππ
β πβπ§ β
βππ + βππ΅
ππ
(Sinnott, 2005) (3.8.7)
This calculates a value of -92 J.kg-1
, meaning this much energy needs to be expended to move 1 kg of
material through the pump. As the flow rate through the pump is 336 kg.s-1
, this would give a total
power output of 31 kW.
38. Page | 33
3.9. Valves
A number of different valves are being used across section four. A brief justification for their use will
be given in this section and a basic specification will also be given for the manual knife gate valve MV-
20/03, found on stream 20.
Control valves are used to control parameters such as flow rate by partially
or fully opening or closing in response to signals received from the
controllers that itβs connected to. This opening or closing is done
automatically by electrical actuators that operate with 4-20mA signals,
which is standard for industry (Sinnott, 2005). There are several types of
control valves available, the type used in section four is called a diaphragm
valve.
Diaphragm valves, shown in figure 3.5, can be broken into two main
categories; saddle type or seat type. Seat type valves will be used in section
four as they are more suited for use in slurry applications as they reduce
blocking issues. Diaphragm valves are also very reliable due to their low
number of moving parts and they can also control flow rates to high
degrees of accuracy (Nesbitt, 2011).
Butterfly valves will be used to regulate the flow of air into each flotation cell. The closing mechanism
in a butterfly valve takes the form of a disk. They are a popular choice of valve due to their low cost
and light weight (Dickenson, 1999).
Knife gate valves will be used for the manual valves. A gate valve is a valve that opens by lifting a
round gate out of the path of the fluid. Knife gate valves have the ability to cut through thick liquids
in slurries making them ideal for this process (Nesbitt, 2011). However, they should not be used to
regulate flow unless they have been specifically designed for that purpose, for that reason they are
only used as isolation valves in this process. All valves used in section four will be constructed from
SS-316, with the butterfly valves being the only exceptions, as they will be constructed from mild steel.
To specify a valve the valve coefficient must be calculated. This can be done using equation 3.8.1:
πΆ π = πβ
πΊ
βπ
(Sinnott, 2005) (3.8.1)
Where πΆ π is the valve coefficient, π is the flow rate in gallons per minute, πΊ is the specific gravity of
the slurry and βπ is the pressure drop in psi. The values for which can be seen in table 3.6 below.
Table 3.6: Values for calculating valve coefficient.
Parameters Units Value
π Flow Rate gallons/min 3140
πΊ Specific Gravity - 1.48
βπ Pressure Drop psi 7.1
πΆ π Valve Coefficient - 1433
Figure 3.5: Diaphragm valve
(Sinnott, 2005).
