Gravity concentration presentation

1. GRAVITY CONCENTRATION
 Gravity concentration methods separate
minerals of different specific gravity.
 They are used to treat a great variaty of
materials [ranging from Au ( sp. gr. 19.3 ) to
coal ( sp. gr. 1.3 ) ]
 Limitations for application :
1. If the sp. gr. difference between the minerals
is less
2. If the liberation is achieved in fine sizes.
3. If high capacity is needed especially in finer
sizes.
4. Low grade and complex ores.
1

2.  Gravity concentration methods remained, however the main
concentrating methods for iron , tungsten, tin ores and coal.
 Gravity methods are usually preferred to flotation due to its low
cost . Minerals liberated at sizes above flotation range may be
concentrated even more economically using gravity methods (also
cause efficient dewatering due to decreased surface area.)
2

3.  In recent years, many companies have re – evaluated gravity
systems due to :
1. Increasing cost of flotation reagents
2. Relative simplicity of gravity processes
3. Procedure comparatively little environmental pollution
3

4.  Uses :
1. To produce final concentrates.
a-) Coarsely liberated mineral
b-) Low value minerals
c-) Those which are not suitable to flotation
d-) Plaser deposits.
2. As pre – concentration
3. To recover residual valuable heavy minerals in
flotation tailings.
4

5. Principles of Gravity Concentrations
1. It is essential for effective separation that a marked
density difference exists between the mineral and
the gangue.
In order to examine the amenability of
concentration of certain mineral by gravity
concentration methods, Concentration Criteria is
used.
5

6.   H : Sp. gr. of heavy mineral
 L : Sp. gr. of light mineral
 F : Sp. gr. of fluid
6
F
L
F






 H
Criteria
ion
Concentrat

7.  If conc. crit. > 3 Gravity sep. is easy in all sizes.
 If conc. crit. > 2 No difficulty, effective concentration
is possible down to the size of fine sands. Clean
concentrate is produced, but it is difficult to obtain
clean tailing. The tonnage of middling is large.
 If conc. crit. = 2.5 – 1.75 Commercial separation is
possible down to 100 mesh
 If conc. crit. = 1.75 – 1.5 The limit of fineness is
around 10 mesh.
 If conc. crit.  1.25 Gravity separation is not
commercially feasible .The separation is very difficult
even impossible.
7

8. 2. The motion of a particle in a fluid is
dependent not only specific gravity , but
also on its size ( and shape ) . Larger
particles will be affected more than smaller
ones.
The efficiency of gravity processes
increases with particle size and the particles
should be sufficiently coarse to move in
accordance with Newton’s Law. ( coarse
particles overcome surface friction during
their movement.)
8

9. 3. Close size control of feeds to gravity processes is required in
order to reduce the size effect and make the relative motion
of the particles specific gravity dependent .(The feed to jigs,
cones, spirals should be screened while in shaking table
utilization, classified feed is fed.)
9

10. 4. It is common practice in most gravity
concentrators to remove particles < 10 m from
the feed, because they are extremely sensitive to
the presence of slime.
5. To minimize degradation of friable minerals, it
should be made reduction of slurry pumping (as
much use of gravitational flow as possible)
6. Correct water balance in gravity circuits is
essential (optimum feed pulp density )
Nucleonic density gauges control the water
addition to the new feed.
10

11. 7. If the ore contains appreciable amount of
sulphide minerals :
a-) If the primary grind is finer than about
300 m, these should be removed by
flotation prior to gravity concentration ( as
they reduce the performance of spirals,
tables etc.)
b-) If the primary grind is too coarse for
effective sulphide flotation, then the gravity
concentrate must be reground prior to
removal of the sulphides
11

12.  Size Ranges of Feed in Various Gravity
Separation Applications.
 DMS : Down to 3 mm ( sometimes down to
0.5 mm if centrifuge is applied )
 Jigs : Down to 150 m ( or 75 m ). In both ,
top size  250 mm
 Sluice : 25 mm – 0.25 mm
 Reichert Cone : 3 mm – 30 m ( normal range
100 – 600 m )
 Spiral : 3 mm – 75 m
 Shaking Table : 3 mm – 25 m ( sand table >
100 m )
 Tilting Table : 100 – 5 m
12

13. HEAVY MEDIUM SEPARATION (HMS) (or
Dense Medium Separation – DMS - )
 HMS is the simplest of all gravity processes and
has long been a standard laboratory method for
separating minerals of different specific gravity.
13

14.  Industrial Uses :
1. Pre – concentration of minerals ( For metalliferous ores rejection
of gangue prior to final liberation )
2. In coal preparation ( to produce final clean coal )
14

15.  Advantages :
1. It has ability to make sharp separation at
any required density .
2. It has a high degree of efficiency , even in
the presence of high percentage of near –
density material.
 The process is, however, rather expensive ,
mainly due to the equipment needed for
the regeneration of the medium.
15

16.  The process is mostly used when the density
difference occurs at a coarse particle size (
Efficiency decreases with the size due to the
slower rate of settling of the particles.) Particles
should be larger than about 3 mm in diameter, in
which case separation can be effective on a
difference in specific gravity of 0.1 or less. There is
no upper size limit.
 Separation down to 500 m, and less, in size can
be made by the use of centrifugal separators.
16

17.  HMS is possible with ores in which the minerals are
coarsely aggregated. ( If the values are finely
disseminated throughout the host rock, suitable
density difference between crushed particles
cannot be developed.)
 Heavy liquids are used in the laboratory HMS.
Thick suspensions of fine solids (pulp ) are used in
industrial applications.
17

18. THE HEAVY MEDIUM
 LIQUIDS:
 Heavy liquid testing ( Sink – and – Float Process ) may
be performed in the laboratory :
1. To determine the feasibility of HMS on a particular
ore.
2. To determine the economic separating density
3. To assess the efficiency of an existing HM circuit (
performance test )
Inorganic salts (ZnCl2, CaCl2)
Liquids used in HMS
Organic liquids
18

19. 19
Material Chemical Formula Max
sp. gr.
Calcium Chloride CaCl2 1,30
Zinc Chloride ZnCl2 2,07
Carbon Tetra Chloride CCl4 1,59
Methylene bromide CH2 Br2 2,96
Bromoform CH Br3 2,89
Clerici solution CH2 (COOTl )2+HCOOTl 4,20
Tetrabrom Ethane CH2 Br . CBr3 2,96
Sodium Polytungstate 3,1

20.  Aqueous solutions of Na - polytungstate Density
up to 3.1 (Non – volatile , non – toxic)
 Clerici solution (Thallium formate – Thallium
malonate solution)
Density up to 4.2 – Exteremely poisonous
 By the use of Magneto hydrostatics
Density up to 12 (a paramagnetic salt situated in a
magnetic field gradient)
(used for fine size particles of about 50 µm)
 Many organic liquids give off toxic fumes and
must be used with adequate ventilation. Clerici
liquids are extremely poisonous and must be
handled with extreme care.
20

21. SUSPENSIONS
 In industrial processes, finely ground solids
suspended in water are used as medium.
 If the concentration of fine solids by volume
< 30% , they behave as simple Newtonian fluids,
but > 30 % by volume the suspension becomes
Non – Newtonian and a certain minimum stress
or yield stress has to be applied before shear will
occur, and the movement of a particle can
commence.
 The shearing force may be increased by applying
centrifugal force. The viscous effect of
suspension may be decreased by agitating the
medium (both will decrease the rigidity of
medium )
21

22.  Properties of solids used to produce a stable
suspension :
1. Sufficiently high density
2. Reasonably low viscosity
3. Must be hard with no tendency to slime
4. Must be readily removed from the surfaces by
washing
5. Must be easily and cheaply recoverable
6. Must resist to chemical attack (corrosion)
7. Must not be affected by the constituent of the
ore.
22

23. 23
Material Max. sp. gr. İn solution Hardness
Sandstone,
quartzite
1.58 5 – 7
Barite 2.05 3 – 3.5
Pyrite 2.38 6 – 6.5
Magnetite 2.50 5.5 – 6.5
Galena 4.00 2.5 – 2.75
Ground Fe - Si 3.40 7.3 – 7.6

24. Galena – Regeneration by flotation
Magnetite – Regeneration by magnetic
separation
Fe – Si - Regeneration by magnetic separation
 Ferrosilicon is an alloy of iron and silicon which
should contain not less than 82 % Fe and 15
– 16 % S.
 If the Si content < 15 %, the alloy will tend to
corrode
 If the Si content > 15 %, the magnetic
susceptibility will be reduced. ( Total losses of
Fe – Si due to losses in the regeneration circuit as
well as corrosion is  0.1 – 2.5 kg / tonne.
24

25. SEPARATING VESSELS
1. Gravitational (Static baths)
2. Centrifugal (Dynamic)
 In both de – slimed feed is used.
25

26. GRAVITATIONAL VESSELS (Cones, Drums,
Baths)
 Here, feed and medium are introduced into the vessel by free
fall.
 Floats are removed by paddles or by overflow.
26

27. WEMCO CONE SEPARATOR 27

28.  Sinks are removed by pump or by external or internal air lift.
 Air lift : The sink drops to the bottom of the cone where it is
picked up by central air –lift which raises the sink to the level of
the surface of the medium and discharges it into the sink
launder.
 Wemco cone separator is widely used for ore treatment since it
has a relatively high sinks capacity.
28

29.  Cone diameter : Up to 6 m
 Feed size : Up to 10 cm
 Capacity : Up to 500 tph
29
Gentle agitation
by rakes to keep
the medium in
suspension.

30. Drum Separators
30
Single compartment

31. 31
Double compartment

32.  Removal of the sink product through the
action of lifters. The lifters empty into the
sink launder.
Diameter : Up to 4.3 m
Lenght : 6 m
Max. capacity : 450 tph
Feed size : Up to 30 cm (usually from 6 mm
to 30 cm)
32

33.  Advantages :
1. Shallow pool depth in the drum minimises
settling out of the medium particles giving a
uniform gravity throughout the drum.
2. Simple, reliable and low maintenance cost.
 Two – compartment drum separator provides
three products. The lighter medium in the first
compartment separates a pure float product. In
the second compartment true sinks and
middlings are separated.
33

34.  Although drum separators have very large
sink capacities, (more suited to the treatment
of metallic ores which have sink product
normally 60 – 80 % of the feed ), they are
usually used in the coal industry because of
their simplicity and reliability.
34

35. Drewboy Bath 35

36.  It has high floats capacity.
 The sinks are lifted out from the bottom of the
bath by a radial – vaned wheel mounted on an
inclined shaft.
 The medium is fed into the bath
1. At the bottom of the vessel
2. With the raw coal
 Tesca bath : With straight wheel and for high
sinks capacity.
36

37. Norwalt Washer
37

38.  Developed in South Africa to wash the coal.
 Feed is introduced into the center of the annular
separating vessel with stirring arms.
 The floats are carried round by stirrers and
discharged over a weir on the other side of the
vessel.
 The sink is dragged along by scrapers and
discharged into an elevator (wheel type).
38

39. CENTRIFUGAL SEPARATORS
 DSM cyclones
 Vorsyl separator
 Larcodems
 Dyna Whirlpool
 Tri – flo separator
39

40. HEAVY MEDIUM CYCLONES
 Used both to clean ores and coals.
 They provide high centrifugal force and low
viscosity in the medium, enabling much finer
separations to be achieved than in
gravitational separators.
 Feed is deslimed at about 0.5 mm (in recent
years, the feed size is lowered as fine as 0.1
mm ) . In United States, HMS with no
desliming has been performed.
40

41. DSM cyclone (Dutch State Mines)
 Used to treat ores and coals in the size range
40 – 0.5 mm.
 The ore is suspended in a very fine medium of
Fe – Si or magnetite.
 The sink product leaves the cyclone in the
apex, while the floats product is discharged via
the central vortex finder.
41

42.  The design differs from conventional dense
medium cyclone or a classification hydrocyclone:
 The cone is much larger (up to 120), and the
vortex finder is much longer, in some instances
extending into the conical region.
 It is thought that fine particles of high and
intermediate density collect and recirculate in the
conical section, forming an autogenous dense
medium through which only particles of greater
density can pass.
42
Water – only – cyclone
( Autogenous dense medium cyclone)

43.  The use of water only cyclones is becoming
widespread in cleaning of coal in the 600 – 150 µm
size range.
 Advantages :
1. Lower capital cost
2. Lower operating cost (as the elimination of medium)
(no regeneration)
43

44. VORSYL SEPARATOR 44

45.  Used for the treatment of small coal sizes up to
about 50 mm. Capacity is up to 120 tph.
 De – slimed raw coal with medium (magnetite) is
introduced tangentially.
 Material of specific gravity < medium passes into
the clean coal outlet.
 Near gravity material and heavy shale particles
move to the wall of the vessel due to the
centrifugal acceleration induced.
45

46.  Heavy particles move in a spiral path down
the chamber towards the base of the vessel
where the drag caused by the proximity of the
orifice plate reduces the tangential velocity
and creates a strong inward flow towards the
throat.
 This carries heavy and near gravity material
into the shallow shale chamber which is
connected by a short duct to a second
chamber (vortex tractor) through a tangential
inlet.
 Such an inward spiral flow permits the use of
a large diameter outlet nozzle without the
passing of an excessive quantity of medium.
46

47. DYNA WHIRLPOOL SEPARATOR
47

48.  Used for treating fine coal as well as minerals
(diamond, fluorspar, tin and lead-zinc ores).
 Size of feed : 0.5 – 30 mm
 Capacity : 100 tph
48

49.  Unit is operated in an inclined position.
 Medium is pumped under pressure into the
lower end. Rotating medium creates a vortex.
 Float particles passes down the vortex and
does not contact the outer walls of the unit,
thus reducing wear.
 Heavy sink particles of the feed penetrate the
rising medium towards the outher wall of the
unit and discharged through the sink
discharge pipe.
49

50.  Since the sink discharge is close to the feed
inlet, the sinks are removed from the unit
immediately, again reducing wear
considerably. (Length of flexible hose
connected to sink pipe is used to control back
pressure to control the cut-point.
 Advantages :
1. Less wear, less maintenance cost
2. Lower operating cost (since only medium is
pumped)
3. Can accept large fluctuations in sink / float
rations.
 Dyna Whirlpool Separator has a much higher
sinks capacity.
50

51. LARCODEMS (Large Coal
Dense Medium Separator)
51

52.  Develop for raw coal up to 100 mm in one
vessel.
 The unit consists of a cyclindrical chamber
which is inclined at approximately 30 to the
horizontal.
Diameter : 1.2 m
Length : 3 m
Capacity : 250 tph
Feed : 100 mm – 0.5 mm
52

53.  Medium is fed (introduced) under pressure at
the lower end, feed is fed at the top end.
 At the top end, another tangential outlet
connected to vortex tractor.
 Clean coal after separation is removed
through the bottom outlet.
 High specific gravity particles pass rapidly to
the separator wall and are removed through
vortex tractor.
53

54. TYPICAL DYNA WHIRLPOOL PROCESS FLOW CIRCUIT
(Regeneration of Magnetic Medium)
54

55.  Fe – Si losses can account for 10 -35 % of the
total operating cost of HMS plant.
 Magnetic losses  0,5 -1,5 kg / t
55

56. TRI – FLO SEPARATOR 56

57.  Can be regarded as two Dyna Whirlpool
Separators joined in series.
 Used for coal, metalliferous and non-metallic
ores.
 Tri – flo separator can be operated with two
media of differing densities, or single medium
density for two- stage treatment.
 Single medium with two stage treatment
(scavenger). Float of the first unit is treated
again to remove clean float (in metalliferous).
57

58.  In Italy, by using 2 different density medium,
it treats +1 mm fluorspar – galena ore.
 High sp. gr. medium : 3.2
Low sp. gr. medium : 2.75
 Sink 1 : 41.5 % Pb
40 % CaF2
RPb : 90.3 %
58
Sink 2 : 91.8 % CaF2
0.46 % Pb
RCaF2 : 90 %

59.  Two – stages of separation increase the
sharpness of separation.
 Size of Tri – Flo unit : 250 mm – 500 mm in 
 Capacities : 15 tph – 90 tph
59

60. Laboratory HEAVY Liquid Tests ( Sink
and Float Test)
60

61. 61

62.  Purposes:
1. To determine the feasibility of a particular
ore to HMS ( Washability Curves)
2. Determination of economic separating
density
3. To determine the performance of an existing
plant.
62

63. 63

64. Sp.Gr.
Individual Particles Cumulative Float Cumulative S,ink Ordiin
ate
Axis
a/2+b
Wt.%
a
Ash
%
Ash
Cont.
Cum.
Wt.
%
b
Ash
Cont.t
Aver.
Ash
%
Cum.
Wt6.
%
Ash
Cont.
Aver.
Ash
%
-1,35 8,59 3,31 28,43 8,59 28,43 3,31 100,
0
1481,6
2
14,8
2
4,29
1,35/1,4
0
7,10 4,42 31,38 15,6
9
59,81 3,81 91,4
1
1453,1
9
15,9
0
12,14
1,40/1,4
5
23,73 5,56 131,9
4
39,4
2
191,75 4,86 84,3
1
1421,8
1
16,8
6
27,55
1,45/1,5
0
17,88 7,48 133,7
4
57,3
0
325,49 5,68 60,5
8
1289,8
7
21,2
9
48,36
1,50/1,5
5
16,98 10,87 184,5
7
74,2
8
510,06 6,87 42,7
0
1156,1
3
27,0
8
65,79
1,55/1,6
0
4,45 14,27 63,50 78,7
3
573,56 7,28 25,7
2
971,56 37,7
7
76,50
1,60/1,7
0
2,69 19,08 51,32 81,4
2
624,88 7,67 21,2
7
908,06 42,6
9
80,07
1,70/1,8 2,14 23,99 51,34 83,5 676,22 8,09 18,5 856,74 46,1 82,49
64
1 2 3 4 5 6 7 8 9 10 11

65. 65
Washability Curves of Çanakkale – Çan Lignite
ground to minus 0,5 mm size

66. 1 – Cum. Float Ash(Cum.wt.%float vs.aver. Ash%
of float)
2 – Cum. Sink Ash ( Cum. wt % sink vs. average
ash % of sink)
3 – Elementary Ash (Individual ash % vs. New
ordinate)
4 – Sp. Gr. Curve ( Cum. float wt. vs. sp. gr. )
5 - ± 0,10 sp. gr Distribution
PROBLEM: A tin ore will be precontrated by DMS.The maximum
allowable loss in the tailing is 4 %.
a)What is be the original Sn % of the ore?
b )What will be the separating density?
c) What is the Sn grade of tailing ?
d)What is the Sn content of the preconcentrated feed?
66

67. To determine Economic Separating Density
1 2 3 4 5 6 7
Specific
gravity
fraction
%
Weight
Cumulative
% weight
Assay
(%Sn)
2 * 4 Distribution
(%Sn)
Cumulative
distribution
(%Sn)
-2,55 1,57 1,57 0,03 0,047 0,04 0,04
2,55-2,60 9,22 10,79 0,04 0,369 0,33 0,37
2,60-2,65 26,11 36,90 0,04 1,044 0,93 1,30
2,65-2,70 19,67 56,57 0,04 0,787 0,70 2,00
2,70-2,75 11,91 68,48 0,17 2,025 1,81 3,81
2,75-2,80 10,92 79,40 0,34 3,713 3,32 7,13
2,80-2,85 7,87 87,27 0,37 2,912 2,60 9,73
2,85-2,90 2,55 89,82 1,30 3,315 2,96 12,69
+2,90 10,18 100,00 9,60 97,72
8
87,31 100,00
67
1,11 111,940

68. Solution:
a) The original Sn % of the ore is:111.940/100=1.11 % Sn
b) The separating Density is 2.75 .(The loss of Sn is 3.81 which is lower than
4%)
c) Total content of Sn in the float(tailing)/Total weight of float=Sn grade of
float
4.272/68.48=0.06 % Sn
d)Total content of Sn in the sink(preconcentrated feed)/Total weight % of
sink=Sn % of sink
107.668/31.52= 3.42 % Sn
In a continuously operating process, particles with a high and low specific
gravity in comparison with the medium are least affected. e. i. they will go
either float or sink. Particles with specific gravity approaching that of the
medium may not have time to reach sink (or float) and will be misplaced
into the other product.
 Particles with the same sp. gr. as the medium have an equal chance of
reporting to the sink or float product. Efficiency of separation therefore
varies ranging from 100 % to 50 %.
 The difficulty or ease of separation is dependent on the amount of material
present in the feed which is close to the required density of separation .
The amount of near gravity material present is the weight of material in
±0,10 range of the separating density.
68

69. PARTITION – TROMP CURVE
 The efficiency of a particular separating
process depends on its ability to separate
material of sp. gr. close to the medium.
 The efficiency of separations can be
represented by the slope of a Partition, or
Tromp Curve, whatever the quality of feed,
can be used to estimate the performance.
69

70. PARTITION COEFFICIENT
 The percentage of the feed material of a particular
sp gr which reports to the sink.
 Effective density of separation : Density at which
50 % of the particles report to the sink. 70

71.  The probable error of separation = Ecart probable (EP)
 The lower the EP , the nearer to vertical is the line
between 25 % and 75 %, and the more efficient is the
separation.
 An ideal separation , EP = 0
In practice , EP = 0,02 – 0,08
2
or,
2
25
75
B
A
E
d
d
E
P
P




71

72. 50
25
75
d
2
d
d
on
Imperfecti
I



72
 EP is commonly used as a method of assessing
the efficiency of separation in units such as HMS
process but not as spirals, tables etc. due to the
many operating variables which can affect the
separation efficiency. (HMS is simple and
reproducible ).

73.  The partition curve can be used to predict the products that
would be obtained if the feed or separation gravity were
changed. The curves are specific to the vessel for which
they were established and are not affected by the type of
material fed to it, provided:
A)The feed size range is the same(efficiency drops with
size).Below 10mm,centrifugal separators are better than
baths.
B)The separating gravity is in appr the same range. The
higher the effective separation density, the greater the
probable error, due to the increased medium viscosity. EP is
directly proportional to the separation density, all other
factors being the same.
C)Feed rate is the same.
PROBLEM: Find the partition coefficient(performance) of a
DMS devise whose float is the 82.60 % of the feed (by
weight).
73

74. Construction of Partition Curves
Specific
gravity
fraction
(1)
Floats
analysis
(wt % )
(2)
Sink
analysis
(wt %)
(3)
Floats
% of
feed
(4)
Sinks
% of
feed
(5)
Reconstitute
d feed (%)
(6)
Nominal
sp gr
(7)
Partition
coefficient
(4/5 x 100
%)
-1,30 83,34 18,15 68,83 3,15 71,98 - 4,39
1,30-
1,40
10,50 10,82 8,67 1,89 10,56 1,35 17,80
1,40-
1,50
3,35 9,64 2,77 1,68 4,45 1,45 37,75
1,50-
1,60
1,79 13,33 1,48 2,32 3,80 1,55 61,05
1,60-
1,70
0,30 8,37 0,25 1,46 1,71 1,65 85,38
1,70-
1,80
0,16 5,85 0,13 1,02 1,15 1,75 88,70
1,80-
1,90
0,07 5,05 0,06 0,88 0,94 1,85 93,62
1,90- 0,07 4,34 0,06 0,75 0,81 1,95 92,68
74

75.  The probable error of separation of an
operating heavy medium vessel can be
determined by sampling the sink and float
products and performing (float and sink) test
to determine the amount of misplaced
material in each of the products.
75

76. Scale of values of Near-Gravity Material
(Difficulty of the separation) 76
Wt. %
within ± 0,1
gravity of
separation
Separatio
n Problem
Gravity Process
Recommended
Type
0-7 Simple Almost any process Jigs, tables,
spirals
7-10 Moderatly
difficult
Efficient process Sluices,
cones, HMS
10-15 Difficult Efficient process with
good operating
15-25 Very
difficult
Very efficient process
with expert
operation
HMS
Above 25 Formidable Limited to a few
exceptionally
efficient processes
HMS with
close control

77.  e. g. If the coal is produced with an operating
yield of 51 % at an ash content of 12 %, and if
the theoretical yield at this ash content is 55 %,
the organic efficiency is equal to :
%
7
,
92
55
51

77
yield
l
Theoretica
yield
Actual
Efficiency
Organic 

78. 78

79. 79

80. Sp gr
fraction
(1) Float
analysis
(wt %)
(2) Sink
analysis
(wt %)
(3)
Floats
% of
feed
(4)
Sinks
% of
feed
(5)
Reconstitute
d feed (%)
(6)
Partition
coefficient
(%)
(7)
Nominal
sp gr
-1,30 38,84 - 16,24 - 16,24 0 -
1,30-1,35 35,44 0,21 14,82 0,12 14,94 0,8 1,325
1,35-1,40 15,12 0,48 6,32 0,28 6,60 4,2 1,375
1,40-1,45 7,03 1,44 2,94 0,84 3,78 22,2 1,425
1,45-1,50 2,68 3,19 1,12 1,86 2,98 62,41 1,475
1,50-1,55 0,34 2,71 0,14 1,58 1,72 91,86 1,525
1,55-1,60 0,33 5,82 0,14 3,39 3,53 96,03 1,575
1,65-1,70 0,07 2,35 0,03 1,37 1,40 97,85 1,675
1,17-1,75 0,09 3,60 0,04 2,09 2,13 98,12 1,725
+1,75 - 78,12 - 45,4
4
45,44 100 -
Total 100 100 41,852 58,1
8
100
80

81.  Ash % Feed : 47,51 %
Ash % Sink : 75,19 %
Ash % Float : 14,01 %
Feed ( wt %) = Float (%) + Sink (%)
Ash content of feed = Ash content of float + Ash content of sink
100 x 47,51 = Float x 14,01 + (100 – float) x 75,19
Float :41,82 %
Sink : 58,18 %
81

82. JIGGING
 The jig is normally used to concentrate relatively
coarse material. Good separation is possible if
the feed is closely sized (e.g. 3 – 10) and if the sp.
gr. difference is large.
 Many large jig circuits are still operated in the
coal, cassiteriite, tungsten, gold, barytes and iron-
ore industries.
 Jigs have a relatively high unit capacity and can
achieve good recovery of values down to 150 µm
and acceptable recoveries often down to 75 µm.
(High proportions of fines interfere with
performance.)
82

83.  Jigging is the stratification of minerals of
different sp. gr.
 The separation is accomplished in a bed which is
rendered fluid by a pulsating current of water so
as to produce stratification.
 The aim is to dilate the bed of material being
treated and to control the dilation so that the
heavier, small particles penetrate the interstices
of the bed and the larger high sp gr particles fall
under a condition probably similar to hindered
settling.
83

84.  On the pulsion stroke, the bed is normally lifted as a mass , the
bottom particles falling first until the whole bed is loosened.
 On the suction stroke, it then closes slowly again and this is
repeated at every stroke; the frequency varying between 55 –
330 cycle / min.
84

85.  The fine particles tend to pass through the
interstices after the large ones have
become immobile. The motion can be
obtained either by using a fixed sieve jig
and pulsating the water, or by employing a
moving sieve in a simple hand- jig.
85

86. The Jigging Action
 The motion of a particle settling in a viscous
fluid is :
R
g
m
g
m
dt
dv
m 




 '
movement
particle
the
to
due
resistance
Fluid
R
R
-
ac.
gr.
mass
-
ac.
gr.
mass
on
accelerati
mass fluid
displaced
solid





86

87.  At the beginning of the particle movement, since the
velocity is very small, R can be ignored as it is a
function of velocity.
 Volume of particle and displaced fluid are of equal
volume.
g
m
m
m
dt
dv








 

'
fluid
gr.
Sp.
solid
gr.
Sp.
1
f
s





















 








s
s
f
s
f
g
g
dt
dv
87

88.  The initial acceleration of the particle is
independent of size and dependent only on
the densities of the solid and the fluid.
 If the duration of fall is short enough, the
total distance travelled by the particles will be
affected more by the differential initial
acceleration, and therefore by density than by
their terminal velocities, and therefore by size.
 To separate small heavy mineral particles from
large light particles, a short jigging cycle is
necessary.
88

89.  More control and better stratification can be achieved by using
longer, slower strokes, especially with the coarser particle sizes.
 It is good practice to screen the feed to jigs into different size
ranges and to treat them separately.
89

90.  If the mineral particles are examined after a
longer time, they will have attained their
terminal velocities and will be moving at a rate
dependent on their sp gr and size.
 Since the bed is really a very thick suspension of
high density (loosely packed mass with
interstitial water), hindered settling conditions
prevail. 90

91.  The upward flow can be adjusted so that it
overcomes the downward velocity of the fine
light particles and carries them away, thus
achieving separation .
 Hindered settling has a marked effect on the
separation of coarse minerals, for which longer,
slower strokes should be used.
91

92.  At the end of pulsion stroke , as the bed begins
to compact, the larger particles interlock,
allowing the smaller grains to move
downwards through the interstices under the
influence of gravity (consolidation trickling).
92

93. 93

94.  In the jig , the movement of the piston is a
harmonic waveform
94

95.  The speed of flow through bed during jig
cycle is sine – curve.
95

96.  Point A : Upward flow increases, the bed is
forced to open or dilate.
B : Hindered settling phase (particles
move according to mass)
C : Fine grains are pushed upwards
by the flow
D : The coarser grains will fall back
E : Bed will be compacted.
Consolidation trickling now occur.
96

97.  Severe compaction of the bed can be
reduced by the addition of hutch water
(which creates a constant upward flow
through the bed). Thus, suction is reduced by
hutch water.
97

98. TYPES of JIG
 Modern jigs employ a stationary screen
and pulse the water flow through it.
 The significant variations between the
various types of modern jig therefore
become :
1. The method used to cause the pulsation
2. The method of withdrawing the dense
mineral product.
98

99.  The tank or hutch is divided into 2 main
sections: One containing the screen, the other
where the fluid pulse is generated.
 In Denver : Diaphragm generate the water
pulsions
In Harz : Plunger generate pulsions
In Baum : Air pressure generate the water
pulsions.
99

100.  2 – main methods of dense mineral removal
can be employed :
1. Over the screen jigging : Dense mineral
discharges under a suitable weir, while the
lighter mineral overflows a different weir.
2. Through the screen jigging : Dense mineral
particles fall into the hutch where they are
withdrawn by a spiral or bucket elevator.
10
0

101. HARZ JIG (Over the screen type)
 Oldest type. Plunger moves up and down vertically in a
separate compartment.
 Four successive compartments are placed in series.
 A high grade concentrate is produced in the 1.
compartment
 Tailing is overflowing the final compartment.
101

102. DENVER MINERAL JIG (Through the
screen type)
102

103.  Mostly used for removing heavy minerals
from closed grinding circuits, thus preventing
overgrinding.
 The rotary valve can be adjusted so as to
open at any desired part of the jig cycle
(synchronisation between the valve and the
diaphragm)
 By suitable adjustment of the valve, even
complete neutralisation of the suction stroke
with hydraulic water can be achieved.
10
3

104.  Ragging : Coarse, heavy particles placed on
the jig screen if dense minerals have smaller
sizes than the aperture of the screen.
 Ragging mineral can be denser or less denser
than the dense mineral.
 Generally, the ragging material is evenly sized.
The feed flows across the ragging. Grains with
high sp gr penetrate through the ragging and
screen and is drawn off as concentrate from
the hutch.
10
4

105. CIRCULAR or RADIAL JIG
 It is designed (developed) to compensate for the
increase in cross – flow velocity over the jig bed,
caused by the addition of hutch water.
 Feed is introduced in the centre and flows radially
over the jig bed towards the tailing discharge at the
circumference.
10
5

106. 106

107.  Example : Cleaveland – IHC Jig (For minerals)
 Design features are :
1. It is a circular jig in plain view
2. The jig compartment are trapezoidal to reduce the
effect of excessive top water.
3. Radial arrangement of the jig compartment saves
space, provides for simplified feed distribution and
allows the use of a raking device which levels the jig
bed surface as it rotates.
4. Has very large capacity (For 7,5 m diameter, 10
tph/m2 or 41,7 m2 gives 417 tph). Feed size : Max. 25
mm
10
7

108.  In Malaysia and Thailand, it is used for the
treatment of Au, diamonds, iron ore etc.
 In IHC and Denver Jigs, the harmonic motion of
the conventional jig piston is replaced by an
asymmetrical saw – tooth movement of the
diaphragm, with a rapid- short upward, followed
by a slow- long downward stroke.
10
8

109.  This gives the finer particles more time to settle in
the bed, thus reducing their loss.(Jig being capable
of accepting particles as fine as 60 µm.
109

110. Remer Jig
 All four
compartments are
actuated by a
common
mechanism.
(Moving hutch)
110

111. COAL JIGS (Air Pulsated)
 Coal jigs are preferred to the more expensive
Heavy Medium Process when the coal has
relatively little near – gravity material.
11
1

112. 112

113. Baum Jig
11
3

114.  By using air , more suitable pulsation cycles can be
achieved.
 To supplement the pulsion action and maintain
the bed in an open state for a longer time,
additional water is fed to the hutch when the bed
is settling (the downward or suction stroke).
 It consists of a sloping screen surface. A number of
pockets occur along the surface for the removal of
dense material. The refuse discharge is
automatically controlled (Automatic refuse
extractor). It consists of a float immersed in the
bed.
 An increase in the depth of refuse raises the float,
which automatically controls the refuse discharge,
by adjusting the height of the moving gate.
11
4

115.  Baum jig can handle large tonnages up to
1000 tph, with a wide size range. However,
the distribution of the stratification force,
being on one side of the jig , tends to cause
unequal force along the width of the jig
screen.
 Uneven stratification due to the unequal
stratification force causes some loss in the
efficiency of separation of the coal from its
heavier impurities.
11
5

116. Batac Jig 11
6

117.  It has no side air chamber like the Baum Jig. It
is designed with a series of multiple air
chambers, usually two to a cell, extending
under the jig screen for its full width, thus
giving uniform air distribution.
 The jig uses electronically controlled air valves
which provide sharp cut – off of the air input
and exhaust.Both inlet and outlet valves are
indefinetly variable with regard to the length
of stroke so that proper stratification may be
achieved. Therefore, Batac Jig can wash both
coarse and fine sizes well.
11
7

118.  As, there is no adjacent chamber, it can be
designed with wider beds that permit higher
throughput in a given space.
 Raw coal is fed across the full width of the jig
bed No: 1. Pulsation in the first 2 – cell
separate coal and middlings from coarse
refuse. The layer of heavy refuse serves as a jig
bed.The bed depth of the two – cell is adjusted
automatically to open discharge gate.
11
8

119.  Cells 3 and 4 have feldspar beds to carry coal,
middlings and fine refuse from cell 2. The fine
refuse sifts down through the bed and screen.
Cell 5 operates like cell 2, with coarse midlings
serving as the jig bed and than discharges at
the end of this cell.
 Jig normally operates at 60 impulses / minute.
The jigging air required is at 6,5 psi pressure.
 Finer middlings and clean coal pass into the cell
6 which has feldspar bed.Fine middlings sift
down through the bed and screen, leaving
clean coal to overflow final cell.
11
9

120. SLUICES – PINCHED SLUICES
 The Sluices Box : is an
inclined trough through
which feed is washed after
removing large pieces by
means of grizzly or
trommel.
 Riffles are placed in the
bottom in order to create
bed turbulence, establish a
hindered settling zone.
120

121.  Operation is intermittent to remove concentrate and other
products.(Starting from the end, the content is raked against the
flow and riffles are gradually removed to remove concentrate).
12
1

122. 12
2
Pinched Sluices
It is an inclined trough about 1m long, 20
cm in width at the feed end, and 2,5 cm at
the discharge. Concentrate is discharged
continuously. No riffle at the bottom.

123.  Pulp density : 50 – 60 % solids by weight.
 Feed size : 10 mesh (1,68 mm) - 37µm.
 Throughput : 4 – 10 tph depending on particle size.
 Slope :  15 
12
3

124. REICHERT CONE
 It is a wet pre – concentrating device based
or flowing film.
 Has a high capacity, normally in the range
65 – 90 tph, but in exceptional cases 40 –
100 tph.
 Feed density : 55 – 70 % solids by weight
(Fluctuation in density affects the efficient
of separation)
 Feed size : Up to 3 mm, can treat as fine as
30 µm. Most efficient in the 100 – 600 µm.
12
4

125.  Reichert cone has similar operational
principle to that of pinched sluice, but it
has no wall effect to generate turbulance
(the pulp flow is not restricted by side –
wall effect).
 Cone gives sharper separations than a
pinched sluice.
 Reichert Cone was developed in Australia
to treat titanium bearing beach sands.
12
5

126. 12
6

127.  The cones are made of fiberglass.It is
mounted in a circular frames over 6 m high.
 Each cone is 2 m diameter. There are no
moving part.
 Single unit is consisted of double and single
cones, together with trays (for rougher
concentration, four double – single cone in
series).
12
7

128.  The feed is distributed around the periphery of
the cone.
 Heavy particles separate to the bottom of the
film as the pulp flows towards the centre of the
cone. Concentrate is removed by an annular slot
of the concentrating cone. Tailing is flowing over
the slot.
 Pinched sluices and cones have relatively low
upgrading ratios. The products is retreated in
cleaner and scavenger circuits.
12
8

129. 129

130. 13
0

131.  Reichert Cone is used for preconcentration of
tin, gold, tungsten, magnetite.
 Cones, due to their high capacities and low
operating costs, have replaced spirals and
shaking tables.
 In South Africa, 68 Reichert Cones treat 34000
tph to treat flotation tailings.The feed is
upgraded of about 200:1, then is further
cleaned on shaking tables to produce final
concentrate.
13
1

132. HUMPHREYS SPIRAL 13
2

133.  Used first chromite concentration.
 Most extensive usage has been in the
treatment of heavy mineral sand deposits
such as ilmenite, rutile, zircone, monazite.
 It is composed of a helical conduit of
modified semi-circular cross-section.
 Feed pulp : 15 – 45 % solids by weight.
 Feed size : 3 mm – 75 µm.
13
3

134.  The particles stratify due to the combined
effect of ;
1. Centrifugal force
2. Differential settling rates of particles
3. The effect of interstitial trickling through the
flowing particle bed.
 Hindered settling is also effective.
13
4

135.  Ports for the removal of heavier particles are
located at the lowest points in the cross – section.
Wash water flows outwardly across the
concentrate band to clean the concentrate.
 The grade of concentrate taken from the
descending ports progressively decreases.
 Tailing is discharged from the lower end of the
spiral.
13
5

136.  By the development of spiral technology,
new spirals are introduced as;
1. With only one concentrate take – off, at the
bottom of the spiral
2. Without added wash water (Lower cost, easy
operating, simple maintenance) used for Au,
Tin.
 To treat fine coal, specifically designed
spirals are introduced ( 0,2 – 1 mm).
13
6

137.  Double – spiral concentrators have two starts around a
common column.
 In Canada, 4000 double start spirals are used to clean
hematite ore at 5000 tph capacity at 90 % recovery.
 Spirals are made with slopes of varying steepness the
angle effecting the specific gravity.
 Shallow angles are used, e.i. to treat coal (1 – 3 tph)
 Steeper angles are used to treat minerals (2 – 6 tph)
 Spiral Length : 5 or more turns for rougher
3 for cleaner
13
7

138. SHAKING TABLES
13
8

139.  It is the most efficient gravity concentrator.
 It is principle of working is based on flowing film.
 When a flowing film of water flows over a flat
inclined surface, the water closest to the bottom is
retarded by the friction of water adsorbed on the
surface.
 If minerals are introduced into the film, small
particles will not move as rapidly as large particles,
as they will be submerged in the slower moving
portion of the film.
 Particles of high specific gravity will move more
slowly than lighter particles, lateral displacement
will be produced.
13
9

140. Feed : Classified feed (coarser light, small dense).
Pulp density : 25 % solids by weight.
 Water is flowing down, across the riffles.
 Table is vibrated longitudinally – Using slow
forward stroke and rapid turn.(mineral particles
crawl along the deck)
Deck : Inclined, with riffles on it, parallel to the
long axis.
Riffles : Tapered from a maximum height on the
feed side. (Vertical stratification behind the riffles)
14
0

141.  Particles are subjected to 2 – forces:
1. Due to the table motion (asymmetrical acceleration
move along the direction of motion)
2. Due to the flowing film of water (flowing film,
hindered settling, consolidation trickling)
 These forces are perpendicular to each other.
 The smaller, denser particles riding the highest
towards the concentrate launder at far end.
 Larger light particles are washed into the tailings
launder.
14
1

142.  Riffles : Run parallel with the long axis of the
table. Tapered from a maximum height on the
feed side, till die out near the opposite side.
 Vertical stratification takes place behind the
riffles, so that the finest and heaviest particles
are at the bottom and the coarsest and
lighest particles are at the top. (the finer sized
and higher density particles are protected
behind the riffles).
 Final concentration takes place at the
unriffled area at the end of the deck.
14
2

143. 14
3

144. Variables of Shaking Table
 Design Variables :
1. Table shape
2. Table surface material ( Table : wood,
Lining : Linoleum, rubber, plastic with a high
coefficient of friction, fiberglass)
3. Shape of riffles
4. Pattern of riffles
5. Feed presentation
14
4

145.  Operating variables :
1. Table tilt (desk slope)
2. Pulp density of feed ( 20 – 25 % solid by wt for ore,
30 – 35 % solid by wt for coal)
3. Feed rate
4. Wash water
5. Position of product splitter.
6- Running speed : Motor speed , Pulley size
Stroke : Toggle or vibrator settling
Length of stroke : 10 – 25 mm
Number of stroke : 240 – 325 srokes / min.
Amplitude : Handwheel
14
5

146.  Particle size plays important role in table
separation.
 The efficiency of separation decreases with the
increase in size.
 Capacity : 2 tph for 1,5 mm sand
(per deck) 1 tph for fine sand
0,5 tph for 100 – 150 µm feed
12,5 tph for -5 mm
(may be up to 15 mm)
14
6
Ore
cleani
ng
Coal cleaning

147.  Double and triple deck units has improved
the area / capacity ratio.
 Fine feed requires a higher speed and shorter
stroke.
 Coarse feed requires a slower speed and
longer stroke.
 Table slope varies with feed size and it is
greatest for the coarsest and highest gravity
feeds.
 Normal end elevations in ore tabling range
from a maximum of 90 mm for a heavy,
coarse sand, to as little as 6 mm for an
extremely fine feed.
14
7

148.  USES:
 Ore concentrating
tables are used
primarily for
 Coal washing
14
8
Tin
Iron
Tungsten
Barium
Titanium
Zirconium
Especially in U.S.A
( 45 million tons of
metallurgical coal)

149.  Since the shaking table effectively separates
coarse light from fine dense particles, it is
common practice to classify the feed.
 Classification is usually performed in multi –
spigot hydrosizers in order to feed as narrow
a size range on to the table.
 Each spigot is fed to a separate shaking table.
 Sand Tables : Riffled tables : Operate on feed
sizes in the range 3 mm to 100 µm.
 Slime Tables : For -100 µm sizes.
14
9

150. Typical Shaking Table Concentrator Flowsheet 15
0

151. Pneumatıc Tables
 Used where water is at a premium (eg in the
upgrading of asbestos, for seed separation; for
heavy mineral sand deposits)
 Pneumatic tables use a throwing motion to
move the feed along a flat riffle deck, and blow
air continuously up through a porous bed.
15
1

152.  The stratification produced is somewhat
different from that of wet tables.
 In wet tables, the particle size increases and
the density decreases from the concentrate
band to the tailing band.
 On air table, both particle size and density
decrease from the top down (similar to
hydraulic classification)
15
2

153. Bartles – Mozley Tables
 Used for the recovery of very small particles.
 Feed size : -100 µm +5 µm (for Au and Pt, down
to 1 µm)
 Feed rates : Over 5 tph
15
3

154.  The flowing – film thickness is  0,5 – 1 mm
(means that for a 100 µm particle, it is 10
times ; for 5 µm paticle, 200 times)
 Separation is based on the fact that a
suspension of particles is subjected to
continuous shear due to the forward motion
of film across the surface, or movement of the
surface. The concentrator uses orbital motion
to develop shear in the bed.
 Bartles – Mozley tables consists of 40
fiberglass decks, each 1,1 m wide by 1,5 m
long, arranged in two sandwiches, each of 20
decks.
15
4

155.  Each deck is smooth and attached to its
neighbours by 12 mm thick plastic formers
which serve as pulp channel.
 The deck assemblies are supported on two
suspension cables at a 1 -3 angle to the
horizontal, within a free – standing steel
framework.
 An integral drive assembly is located between
2 – deck sandwiches.
15
5

156.  Operation:
1. The feed is distributed evenly to all 40 decks
through a piping system, for a period of 35
minutes.
2. The feed is automatically interrupted, with
simultaneous tilting of the table by an air
cylinder to drain tailing, then further tilting
to about 45 for washing cycle.
3. Low – pressure water is fed to remove the
original.
4. The deck is returned automatically to the
original position, wash water valve is shut
off and feed valve opened to commence the
next cycle.
15
6

157.  Bartles – Mozley tables is a semi – continuous
device.It is used for pre – concentration rather than
the production of the final concentrate (roughing
or scavenging device).
 It is initially developed for the concentration of fine
cassiterite but is also used for the recovery of fine
mineral from old mine dumps.
15
7

158. Bartles – Cross Belt Separator
 It was introduced to upgrade Bartles –
Mozley concentrate.
 It consists of a 2,4 m wide endless PVC
belt, whose top surface is inclined from a
central longitudinal ridge towards both
side edges. Belt moves slowly over drive
and head pulley, a rotating out – of
balance weight imparting orbital motion
to the belt. The belt assembly is
suspended freely from the main frame by
4 wires.
15
8

159.  The feed is introduced along the first half of
the length. Heavy particles are deposited on
the belt while light particles (gangue), held in
suspension by the orbital shear, flows down
the belt.
 The concentrate travels sideways with the belt
through cleaning zone where middling
particles are washed down to a middling
launder.
 Finally the clean concentrates are discharged
over the head pulley.
15
9

160.  The unit of measurement of magnetic flux density
(or magnetic induction )(B) is the Tesla (T).
 B = Nm of lines of forces passing through a unit
area of material.
 The magnetising force which induces the lines of
force through a material is called the Field
Intensity (H) = ampere /metre
(1 ampere /metre = 4 x 10 – 7 Tesla)
16
0