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Potash Fertilizer from Biotite

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Abstract and Figures

A process was developed for the utilization of biotite mica, as a source of potassium, for the production of potassium chloride. The investigation was done in two stages. In the first stage, the kinetics of the reaction between biotite and HCl was studied and various kinetic parameters were evaluated. The reaction appears to be pseudounimolecular with the rate-determining step being the replacement of K+ by H+ in the interlayer space. In the second stage of this investigation, a method for extraction and purification of KCl from biotite was developed and subsequently tested on a pilot scale. Here, KCl was recovered from the mica−acid reaction extract, using ethanol. The residue was treated for recovery of Al3+ as the ammonium alum, and Fe2+ was obtained as FeCl2. Biotite (10 kg) yielded 1.4 kg of KCl (99%). The process appears to be well suited for the commercial production of KCl (as well as ammonium alum) from biotite mica.
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Potash Fertilizer from Biotite
Chandrika Varadachari
Department of Agricultural Chemistry and Soil Science, University of Calcutta, 35 Ballygunge Circular Road,
Calcutta 700 019, India
A process was developed for the utilization of biotite mica, as a source of potassium, for the
production of potassium chloride. The investigation was done in two stages. In the first stage,
the kinetics of the reaction between biotite and HCl was studied and various kinetic parameters
were evaluated. The reaction appears to be pseudounimolecular with the rate-determining step
being the replacement of K+by H+in the interlayer space. In the second stage of this
investigation, a method for extraction and purification of KCl from biotite was developed and
subsequently tested on a pilot scale. Here, KCl was recovered from the mica-acid reaction
extract, using ethanol. The residue was treated for recovery of Al3+as the ammonium alum,
and Fe2+was obtained as FeCl2. Biotite (10 kg) yielded 1.4 kg of KCl (99%). The process appears
to be well suited for the commercial production of KCl (as well as ammonium alum) from biotite
1. Introduction
The primary sources of potassium for fertilizers are,
usually, underground deposits of soluble minerals and
potash-bearing brines. Such deposits, however, occur
predominantly in the Northern Hemisphere; about eight
countries in the North and two in the South account
for the potassium produced and consumed by the rest
of the world. This situation is ironic when one considers
that immense deposits of insoluble potassium silicate
minerals occur throughout the world, mainly in the form
of micas and feldspars. Of the two major types of micas,
the muscovites and biotites, the commercial utility of
the latter is practically negligible. A viable means of
extracting soluble potassium from biotite, which has
about 8% K2O, therefore, holds great potential not only
to help several potash-importing countries to attain
some degree of self-sufficiency but also to augment the
total reserves of usable potassium in the world.
Research on the decomposition of silicates for produc-
ing potash fertilizers was initiated in the U.S.A. as a
consequence of Germany’s embargo on the export of
potassium salts during World War I. These early
attempts have been described in a large number of
patents and papers. The processes involve fusion with
alkalis, hydrofluoric acid, limestone, or salts (Harley,
1953; Collins, 1962; Mellor, 1963; Sauchelli, 1967). For
example, feldspars were fused with lime and salt at
about 1000 °C, potash was leached out, and the residue
was used as cement; similarly, Wyomingite was calcined
with lime and salt, and the potassium was volatilized
as KCl and recovered by cooling; greensand was fused
with ferric chloride or nitrate prior to the recovery of
potash salts (Jacob, 1953; Mellor, 1963). However, only
a few of these processes reached the pilot plant stage
and none attained commercial production on a continu-
ing basis (Harley, 1953). Following the discovery of
large deposits of potash salts in the U.S.A. after the war,
the necessity of the aforesaid attempts was drastically
reduced. Besides, as most of the industrially developed
countries, which were advanced in research on fertilizer
technology, had their own reserves of soluble potassium
salts, this subject lost ground and consequently recent
ligerature pertaining to the production of soluble potash
fertilizers from the insoluble silicates is by and large
Till today, none of the earlier processes has been
adopted even in potash-deficient countries because such
processes are, for various reasons, uneconomical. Most
raw materials are expensive and nonrecoverable in the
course of operation; high temperatures (>1000 °C) are
required in one or more stages of manufacture, causing
large fuel consumption; the reagents, under the fusion
conditions usually adopted, are extremely corrosive and
consequently cause difficulties in the choice of materials
for construction and fabrication.
This investigation was, therefore, undertaken with a
view to developing a suitable process, for the extraction
of potassium from biotite, that involves inexpensive raw
materials and moderate temperatures together with
technically simple operating stages. It was initially
developed on a laboratory scale and subsequently
upgraded to a pilot scale.
Preliminary investigations with biotite indicated that
potassium can be readily solubilized by concentrated
acid solutions. This behavior is unlike that of the other
mica mineral, muscovite, which is very difficult to
solubilize except by heating with phosphoric acid (Vara-
dachari, 1992a). In fact this reaction between muscovite
and phosphoric acid has been utilized for the recovery
of potash salts from muscovite wastes (Varadachari,
1992b,c). Although the same reaction can also be
applied to biotite, the ready solubility of biotite in HCl
as well as H2SO4provides easier alternatives. Of the
various inorganic acids tested (i.e., HCl, H2SO4, and H3-
PO4), HCl was observed to be the most effective in terms
of rate of extraction. Moreover, the volatility of HCl is
an added advantage in the succeeding steps involving
evaporation of excess acid. Considering these factors
as well as others, such as price and availability, HCl
was selected as the acid most suited for this purpose. A
complete process for the extraction and purification of
potash salts was developed. This has been presented
here in two parts. The first part consists of fundamental
studies of the kinetics and mechanism of the reaction
between biotite and HCl. In the second part, methods
for purification and recovery of potash salts and results
of pilot-scale trials, are reported.
Present address: Raman Centre for Applied and Interdis-
ciplinary Sciences, 11 Gangapuri, Calcutta 700 093, India.
E-mail: Tel: 91-33-479 1112.
4768 Ind. Eng. Chem. Res. 1997, 36, 4768-4773
S0888-5885(97)00220-0 CCC: $14.00 © 1997 American Chemical Society
2. Kinetics of Potassium Release from Biotite
by Hydrochloric Acid
2.1. Theory. Optimum concentration and amounts
of reactants necessary for extraction, as well as the
model of the reactant vessel and nature of contact or
flow, are all determined by studies on the reaction
kinetics. This is because kinetics provides information
both on the influence of the amount of reactants on the
rate of a reaction and also on the rate-determining step
of a reaction. Thus, in the common rate equation, (dc/
dt))kcn(Glasstone, 1972), the higher the order of
reaction n, the greater the influence of concentration
on the time taken to complete the process.
The rate-determining step for irreversible reaction of
particles with the surrounding fluid may be one or more
of the following processes (Levenspiel, 1975). (i) Film
diffusion control: diffusion of liquid reactant, through
the film surrounding the particle, to the surface of the
particle. (ii) Ash diffusion control: penetration of liquid
reactant, through the blanket of ash, to the surface of
the unreacted core. (iii) Chemical control: reaction of
liquid with solid at the reactant surface.
One method of deducing the rate-determining step of
a fluid-particle reaction is to note the process of reaction
of particles measured in terms of time for complete
conversion, i.e., by plotting (1 -XB) versus (t/τ) where
B is the particle, here biotite, XBis the fraction of
reactant B (i.e., biotite) converted to product at any time
t, and τis the time for complete conversion of reactant
B (Levenspiel, 1975). By studying the nature of this
curve, the rate-determining step can be evaluated.
Moreover, the relation of particle radius with the time
for complete reaction may also be deduced.
2.2. Materials and Methods. Large, good quality
flakes of biotite (from Ajmer, Rajasthan, India) were
hand-picked, washed with twice its volume of dilute
HCl, made chloride free by washing with water, and air-
dried. The flakes were dry ground in an electric grinder
and sieved to obtain the fraction of mesh size 80-150
B.S. The sample was then dialyzed till chloride-free,
dried at 105 °C, and preserved (the sample contains
0.05% adsorbed water). Chemical analysis of the sample
was done according to the scheme of Shapiro and
Brannock (Maxwell, 1968). XRD was recorded using Ni-
filtered Cu KRradiation on a Philips PW 1730 instru-
ment fitted with a Guinier camera. Thermogravimetry
was done on a Gebruder-Netzsch Instrument (No. 404)
at a heating rate of 10 °C/min.
Reactions were carried out at room temperature (25
°C) in borosilicate glass containers. To the biotite was
pipeted in different amounts HCl (11.55 N) of different
dilutions (HCl:H2O, V:V, 100:0, 90:10, 80:20, and 75:
25), and the container was covered with parafilm to
prevent loss by evaporation and allowed to stand
(mechanical stirring was avoided since the nature and
period of stirring greatly influence reaction rates and
the reproducibility of the data is affected). The solutions
were filtered at intervals of 48, 96, 144, 192, 240, and
312 h, washed, and diluted to volume. The K+in
solution was determined by flame photometry. The
exact strength, of the HCl solutions used, was deter-
mined by titration with recrystallized borax (Vogel,
2.3. Results and Discussion. Chemical analysis
of the sample gave the following results, viz., 16.67%
Si, 9.67% Al, 19.24% Fe, 3.35% Ti, 0.23% Ca, 2.60% Mg,
0.26% Na, 7.75% K, 0.05% H2O-(adsorbed water), and
1.20% H2O+(structural water). XRD showed major
reflections at 10.00, 4.60, 3.32, 2.61, 2.43, 2.17, 2.02,
1.53, and 1.305 Å. TG revealed a single weight loss
above 1000 °C that continued up to 1350 °C. Weight
loss at this temperature is due to loss of lattice OH.
Following the expulsion of OH, the biotite structure
persists to about 1100 °C and subsequently is trans-
formed to high-iron magnetic spinel, leucite, and mullite
(Grim, 1968).
Results on the solubilization of K+by HCl of varying
concentrations and volumes are presented in Figure 1.
Figure 1. Solubilization of K+from biotite as a function of time.
Ind. Eng. Chem. Res., Vol. 36, No. 11, 1997 4769
A representative plot of ln(a/a-x) versus t(ais the
total K+released at infinite time, and xis the amount
released after time t) is given in Figure 2. Rate
constants derived from the K+solubilization curves are
given in Table 1.
From the data obtained, the following observations
may be made.
Rates of release of K+from biotite increase with
increasing concentrations of HCl (Table 1). In some
instances, initial rates of release are greater with more
dilute HCl (Figure 1). However, this effect levels off
with increasing time of reaction, so that ultimately
release with more concentrated HCl is faster than with
more dilute solutions.
Reaction periods may be shortened, if necessary, by
heating the reactants. One such example of K+solu-
bilization at 100 °C is shown in Table 2. It may be seen
from this table that 0.5 mL of HCl/100 mg of biotite can
extract all K+from biotite within 90 min when the
reaction is carried out at 100 °C.
Plots of ln(a/a-x) versus tare linear (only some
representative plots have been shown here in Figure 2).
The reaction is therefore of the first order (Glasstone,
1972). However, the lines will not intercept the Y-axis
at zero unlike as in a first-order reaction. A likely
explanation is that the initial reaction between HCl and
biotite is not of the first order but of a higher order. Such
initially higher orders of reaction could be due to release
of exchangeable K+from edges and surfaces. Here, H2O
could also be a reacting component, in addition to HCl,
and this may be the reason for observed higher initial
rates with more dilute HCl, immediately as the acid
comes in contact with biotite. This process of K+release
would no longer be operative when all exchangeable
positions are depleted of K+. In the later stages,
reaction of first order suggests a mechanism of the type
K+-biotite +H+hH+-biotite +K+.
It may be deduced that the above process, unlike
simple exchange, involves first a diffusion of H+into
the interlayer of biotite followed by the release of K+.
This would be the slowest and hence the rate-determin-
ing step of the reaction. Subsequently, the H+probably
migrates into the octahedral and tetrahedral positions
occupied by Al3+,Mg
, and Fe3+, thereby solu-
bilizing these cations.
Since the first-order rate equation is not valid at lower
values of time t, the straight lines in Figure 2 will not
intercept the Y-axis at zero when time t)0. The rate
equation therefore has to be represented in the form y
)kt +c, where y)ln(a/a-x), kis the rate constant,
and cis the intercept on the Y-axis. The values of k
and cfor the various HCl solutions used during reaction
are given in Table 1.
It may be mentioned here that though the reaction
satisfies the first-order equation, it is actually a
pseudounimolecular reaction. The rate of release has
been shown to depend solely on the concentration of K+,
though the reaction mechanism as depicted in the
earlier equation suggests that [H+] is also a determining
factor. However, since the acid is present in large
excess, the amount used up in the course of reaction is
negligible in comparison to the total, so that its con-
centration may be regarded as remaining constant
throughout. Therefore, in the kinetic equation, rate )
k[K+]n[H+]m, this rate is proportional only to the con-
centration of K+since [H+] is almost constant.
A plot of total K+released after 312 h versus mequiv
of HCl/g of biotite (Figure 3) shows that approximately
65 mequiv of HCl is required for the release of all K+
Figure 2. Application of the first-order rate equation to K+
Table 1. Kinetic Parameters of the Biotite-HCl
strength of HCl vol of HCl
(mL/100 mg of biotite) 10-3k(h-1)C
11.55 N, 34.77% HCl 0.375 8.73 8.62
0.500 9.38 1.15
0.625 23.26 0.03
0.750 27.60 -0.02
10.55 N, 33.41% HCl 0.375 7.64 0.87
0.500 12.15 0.82
0.625 14.93 0.77
0.750 15.63 0.88
9.37 N, 30.06% HCl 0.375 6.51 0.84
0.500 8.85 0.83
0.625 10.94 0.77
0.750 16.15 0.55
8.89 N, 28.62% HCl 0.375 5.76 0.79
0.500 10.42 0.50
0.625 14.06 0.33
0.750 16.67 0.25
Table 2. K+Solubilized from Biotite by HCl (8.49 N) at
100 °C
HCl/100 mg of biotite K+solubilized with time (% ot total K+)
mL mequiv 15
min 30
min 45
min 60
min 90
min 120
0.375 3.183 66.67 80.00 83.33 90.00 91.33 96.67
0.500 4.244 70.00 83.33 90.00 93.33 100.00 100.00
0.625 5.304 75.00 86.67 91.67 96.67 100.00 100.00
4770 Ind. Eng. Chem. Res., Vol. 36, No. 11, 1997
In order to determine the model and rate-limiting step
of the reaction, plots of (1 -XB) versus (t/τ) [the symbols
have been explained earlier] were drawn for two rep-
resentative samples (Figure 4). From the nature of the
curve, it appears that the reaction is either ash diffusion
control or chemical reaction control (Levenspiel, 1975).
Since the rate of dissolution of biotite is probably
controlled by the rate of reaction of H+ions with the
ions in the crystal, the appropriate equation (for chemi-
cal reaction control) was chosen and the validity of this
assumption tested by the following methods [Levenspiel,
1975): For particles of constant as well as shrinking
size, a chemical reaction control gives a relation between
fraction converted XBand time tas (t/τ))1-(1 -XB)1/3.
Values of t/τwere plotted against (1 -XB)1/3 (Figure 5).
Straight lines obtained indicate that the reaction is
indeed a chemical reaction control; i.e., the rate-limiting
step is the reaction between HCl and biotite at the solid
3. Extraction and Purification of Potash Salts
from Biotite
3.1. Process Outline. A flow chart for the process
is depicted in Figure 6. A brief theoretical background
of the various stages is as follows: When HCl is reacted
with biotite, solution A is obtained, which contains all
the major cations of biotite, viz., K+,Al
, and Fe2+,in
solution whereas silica is present as an insoluble
residue. The silica is removed by filtration. Solution
A is evaporated to remove excess HCl; solid B thus
obtained contains mainly KCl, AlCl3, FeCl2, some FeCl3,
and HCl as well as precipitated silica. When stirred
with ethanol, the chlorides of iron and aluminum
dissolve, leaving a residue of KCl and silica impurities.
The KCl is purified by washing and crystallization.
Ethanol is recovered by distillation. Aluminum is
separated from iron by precipitation as the ammonium
alum, leaving an acidic FeCl2solution. This process was
initially developed on a laboratory scale and subse-
quently tried out in a small pilot plant.
3.2. Production of KCl on a Pilot Scale. The
laboratory studies provided valuable information on the
nature of the biotite-HCl reaction that was vital for
designing the pilot plant. Results of kinetic studies at
room temperature revealed both the order of the reac-
tion as well as the rate-limiting step. Such information
is a prerequisite for chemical reaction engineering.
Pilot scale studies were done in a rubber-lined reactor
equipped with a stirrer and three silica-sheathed im-
Figure 3. Amount of K+solubilized after 312 h in terms of the
milliequivalent of HCl added.
Figure 4. Nature of the (1 -XB) versus t/τrelation for K+
Figure 5. The chemical reaction-control equation as applied to
Ind. Eng. Chem. Res., Vol. 36, No. 11, 1997 4771
mersion heaters, for the biotite-HCl reaction stage. Due
to the temperature limitation of the rubber lining, the
reaction temperature could not exceed 80 °C. The
reaction was, therefore, studied in the stirred vessel at
80 °C, whereupon it was observed that reaction was
complete in 4 h. (It may be pointed out here that owing
to the difference in the nature of solid-liquid contact
induced by stirring, reaction periods in laboratory scale
studies are quite different from those of the pilot scale.)
Similarly, the purification stages were developed in the
laboratory scale but the quantification of the reactants
was done at the pilot level.
The various stages of the reaction are shown in Figure
6. Commercial HCl (9.4 N, 40 L) was reacted with 10
kg of biotite at 80 °C for 4 h (stage 1). It was then
filtered and washed in a Nutsche filter using polypro-
pylene cloth (stage 2). The residue (stage 2a) contained
silica with titanium dioxide as an impurity. The filtrate
was evaporated in an all-glass distillation unit, under
mild vacuum (stage 3). The residue, after evaporation
(stage 3a), contained crystals of KCl, AlCl3, and FeCl2,
as well as precipitated silica and a small amount of
MgCl2. This was stirred with 25 L of rectified spirit
(90% ethanol) and filtered (stages 4 and 5). The residue
was washed further with rectified spirit (8 L).
The residue obtained on ethanol extraction (stage 5a)
contained KCl together with precipitated silica and
some adsorbed iron and aluminum impurities. This was
dissolved in5Lofwater, and the silica impurities were
filtered out (stage 6). The solution when boiled pro-
duced a precipitate of red iron oxide, which was again
removed by filtration (stage 7). The remaining solution,
when dried, produced crystals of KCl (stages 7b-7d).
Thus, 1.4 kg of KCl of 99% purity was obtained. This
amounts to recovery of about 95% of the total K+in the
The ethanolic extract (stage 5b) contained AlCl3,
FeCl2, HCl, some MgCl2, and H2O. Ethanol, from this
ethanolic extract, was recovered by fractional distillation
(stage 8). The distillate was free of acid; recovery of
ethanol was about 93%. To the residue remaining after
ethanol recovery (stage 8b),3Lofwater was added,
followed by4LofH
SO4(36 N). The solution was
stirred well to dissolve the crystals; then 2.6 L of
Figure 6. Flow chart for production of KCl from biotite.
4772 Ind. Eng. Chem. Res., Vol. 36, No. 11, 1997
ammonia solution (14 N) was added slowly with stirring
(stage 9). Ammonium alum crystallized and was re-
covered by filtration (stage 10). About 13 kg of am-
monium alum was thereby obtained with a purity of
99.9% and an Fe content of 0.005%. The solution
remaining after recovery of alum (stage 10b) contained
mainly FeCl2and HCl together with some MgCl2and
FeCl3. The HCl was removed by evaporation (stage 11)
togive8Lof9Nacid (stage 11d). The remaining soldis
contained 6 kg of FeCl2of 90% purity (stage 11a).
In the various stages of the pilot plant production,
no particularly difficult problems were encountered. The
only stage that was somewhat problematic was the
filtration (stage 2) of the silicia residue from the acidic
extract. The process was quite slow and the filter cloth
tended to clog; mechanical stirring of the residue with
a spatula was, therefore, necessary to facilitate filtra-
tion. Apart from this stage, other processes like distil-
lation and filtration of coarse-grained crystals are
standard operations that could be quite easily carried
Summary and Conclusion
Kinetics of reaction of biotite with HCl revealed that
solubilization of K+from biotite follows first-order
kinetics. It is a pseudounimolecular reaction, and the
rate-determining step is the replacement of K+by H+
in the interlayer space. Various kinetic parameters
were obtained. A method for extraction and recovery
of KCl from biotite was developed in the laboratory and
subsequently tried out on a pilot scale. In this method,
the extract from the biotite-HCl reaction was evapo-
rated and the residue washed with ethanol; precipitated
KCl was purified. Ethanol was recovered from the
extract and remaining solids were treated with H2SO4
and NH4OH to precipitate Al3+as the ammonium alum;
FeCl2was recovered in the filtrate.
In this process, by utilizing low value raw materials,
commercially important products and byproducts are
obtained. The major raw material, biotite, is practically
a waste material with little commercial value; the other
raw material, hydrochloric acid, is very cheap and
widely available. Apart from these, sulfuric acid and
ammonia are also used but these are completely recov-
ered as their byproducts. All the products, viz., potas-
sium chloride, ammonium alum, ferrous chloride, and
amorphous silica, have ample markets as well as
commercial value.
In conclusion, it appears that this process is well
suited for the commercial exploitation of biotite mica
for the production of potash fertilizer. It may be
particularly useful for several countries that lack soluble
potash ores but contain low value biotite deposits; for
others, the recovery of substantial quantities of alumi-
num may provide greater impetus.
The author thanks the Council of Scientific & Indus-
trial Research and the Department of Science & Tech-
nology, Government of India, for financial support. She
is also grateful to Professor S. K. Mukherjee and Dr.
Kunal Ghosh for their advice.
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Received for review March 18, 1997
Revised manuscript received June 25, 1997
Accepted July 1, 1997X
XAbstract published in Advance ACS Abstracts, September
1, 1997.
Ind. Eng. Chem. Res., Vol. 36, No. 11, 1997 4773
... The structure of oxalic acid (C 2 H 2 O 4 ) is composed of two adjacent carboxyl groups, which increases the dissociation constant (pK a1 1.25, pK a2 3.81) (Lide, 2004;Riemenschneider and Tanifuji, 2011) and facilitates the formation of oxalate-metal complexes and the precipitation of metal oxalates, depending on the metal and the chemical conditions (Gadd et al., 2014;Sarver and Brinton, 1927;Sayer and Gadd, 1997). Acidulation of micaceous minerals promotes K release due to the replacement of K + by H + ions in the interlayer space of the mineral (Schimicoscki et al., 2020;Varadachari, 1997). Additionally, the removal of Al, Fe, and Mg cations from the octahedral sheet disturbs the morphology of glauconite, loosening the mineral structure (Hassan and Baioumy, 2006;Schimicoscki et al., 2020). ...
Full-text available
Although Brazil is one of the world's leading exporter of agricultural products, the country is highly dependent on the importation of potassic fertilizers. K-bearing silicate rocks are reported as potential solutions to reduce external K dependency. This work evaluated K extraction from silicate Verdete rock, a glauconite-bearing rock containing 10 % of K2O, by solubilization with organic acids. Firstly, Verdete rock was reacted during 3-120 h with solutions of citric or oxalic acid at 2 % (m/v) in Erlenmeyer flasks by shaking. Oxalic acid extracted 6.5 % of K in Verdete, while citric acid extracted 2.3 %. Another experiment was performed to evaluate the effect of various oxalic acid concentrations (2, 4, 6, 8, and 10 %) and differing reaction times (12, 24, 48, and 72 h) on K extraction from Verdete rock. Soluble K concentration nearly doubled with the increase of reaction time from 12 to 72 h, rising from 20 to 37 mg L–1. Increments in K extraction were obtained by increasing oxalic acid concentrations up to 6 % and above this concentration, no significant gain was observed. The X-ray diffraction data showed that K extraction resulted from the formation of oxalate-metal complexes with metals in Verdete rock.
Silicate rocks (mica, feldspar, glauconite, and sericite) having potash values of 7–11% are evaluated as an alternative potash source. Direct leaching of the feed sample in different lixiviants yielded very less potash extraction (1–5%). Heat treatment with suitable flux such as calcium chloride resulted in 100% potash recovery from mica and glauconite, whereas it was limited to 71% and 65% for feldspar and sericite. The activation energy for CaCl2 reaction with mica, feldspar, glauconite, and sericite is 64.35, 70.14, 69.17 and 86.63 kJ/mol. K – extraction is a function of silica-alumina ratio and association of K – bearing phase.
The present study aims at the recovery of potassium from muscovite mica (which contains K2O; ~10 wt%) using NaCl-roasting coupled with H2SO4-leaching process. The preliminary acid leaching studies applying different mineral acids resulted in a potassium recovery of 8%–18%. The optimum leaching conditions for the maximum recovery were 4 mol/L H2SO4, 60 min leaching time and liquid-solid ratio 4 mL/g at 90 °C. However, the roasting of muscovite with additive NaCl (muscovite: NaCl mass ratio of 1:1.00, 900 °C, 45 min) followed by H2SO4-leaching (95 °C, 60 min) extracted potassium to the tune of 98%. Under similar roasting conditions, the H2O-leaching process extracted only 60% of potassium. The effects of various roasting and leaching parameters such as temperature, time, NaCl concentration, acid concentration, liquid-solid ratio on potassium extraction were evaluated. The appearance of the sylvite (KCl) mineral phase in the NaCl-roasted muscovite and its disappearance in the acid/water leached residue confirmed the physical and chemical distortions of the muscovite crystal structure. The possible mechanism of potassium release from the complex muscovite structure was elucidated based on available literature substantiated by characterizations using X-ray diffraction (XRD) and scanning electron microscopy with energy dispersive X-rays spectroscopy (SEM-EDX).
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The third crucial macronutrient required for the development and growth of plants in addition to nitrogen and phosphorous is potassium. The quantity of soluble potassium in soil that could be directly taken up by plants is less and present in the form of evaporite deposits situated in countries lying in the Northern hemisphere such as Canada, Belarus, Israel, USA and Russia. All other countries are dependent on imports from these countries to fulfil their potash requirement. But a major part of potassium exists in insoluble form as silicate minerals abundantly in these countries which can compensate for the potassium deficit. Mostly feldspar and feldspathoid from the tectosilicate group and micaceous minerals from the phyllosilicate group are the common potassium bearing silicates containing 5–15% K2O. Tectosilicates are framework silicates in which potassium is present in the three-dimensional silicate tetrahedra. Phyllosilicates are sheet silicates having a two-dimensional parallel sheet structure and potassium ions are located in the interlayers of the silicate structure. Various processes have been adopted to release the potassium from these silicate resources such as bio-leaching using various algal, bacterial and yeast strains, acid leaching with different acids of varying concentrations, base leaching with diverse strategies and roast-leach studies with alkali and alkaline salts to optimise the recovery. This review summarizes versatile methodologies that have been approached and the scope of various works that could be further accomplished.
The present article reports the recovery of potassium from mica scrap using carbide lime waste and NaCl. The acid leaching of mica [4 M H2SO4 at 90 °C] can recover a maximum of 25% potassium. However, in the chlorination roasting (with CLW and NaCl) water leaching process, it was possible to recover potassium to the tune of 94% [900 °C, 50 min roasting; mica:CLW:NaCl 1:0.75:1]. The formation of water-soluble phase like sylvite (KCl) and water-insoluble phases like anorthite (CaAl2Si2O8) and nepheline (NaAlSiO4) confirms the complete breakdown of the mica crystal structure. The extraction mechanism proceeds with the formation of a transition state followed by the release of potassium as KCl. The characterisation studies using ICP-OES, XRD, FTIR and FESEM support the proposed extraction mechanism. The salt produced by the evaporation of roast-leached liquor confirms the presence of sylvite and halite only, which can be separated to get the fertiliser grade KCl.Graphic Abstract
The present paper reports the effective utilization of marble sludge powder (MSP) for the recovery of potash values from waste mica scrap using chlorination roasting-water leaching method. Characterization studies indicated the presence of dolomite as the major mineral phase in MSP, whereas muscovite and quartz were observed in the mica sample. The acid leaching studies suggest a maximum of 22% potash recovery under conditions: 4 M H2SO4 acid, particle size of ∼100 µm, stirring speed of 600 r/min, leaching temperature of 75°C, and leaching time of 90 min. The chlorination roasting-water leaching process was adopted to achieve the lowest level of 80%–90% potash recovery. The optimum conditions for the recovery of ∼93% potash from mica (∼8.6wt% K2O) requires 900°C roasting temperature, 30 min roasting time, and 1:1:0.75 mass ratio of mica: MSP: NaCl. The roasting temperature and amount of NaCl are found to be the most important factors for the recovery process. The reaction mechanism suggests the formation of different mineral phases, including sylvite (KCl), wollastonite, kyanite, and enstatite, during roasting, which were confirmed by X-ray diffraction (XRD) analyses and scanning electron microscopy (SEM) morphologies. The MSP-blended NaCl additive is more effective for potash recovery compared with the other reported commercial roasting additives.
Biotite is a potential potassium mineral resource and the dissolution of biotite in H2SO4 solution is a green chemistry approach for recovering potassium. To estimate the efficiency of leaching process and to understand the behavior of leaching reactors, the kinetics and mechanism of leaching potassium from biotite in H2SO4 solution were studied in this work. The leaching experiments show that leached fractions of K2O increase with the increasing of temperatures and H2SO4 concentrations, and yet with the decreasing of biotite particle sizes. According to the shrinking core model, the leached fractions of K2O were fitted better by diffusion through product layer. A kinetics expression of diffusion-controlled model on three parameters was acquired and it is appropriate to predict the leached fractions of K2O from biotite in H2SO4 solution. The dissolution of biotite in H2SO4 solution was interpreted by electrochemical mechanism coupling with characterization of the acid-leached residues. This research results can aid in the design of leaching process for recovering potassium sustainably from biotite.
This review work reports the recovery of potassium from different rocks and mineral sources. The global demand for potassium is rising consistently due to the growth of agricultural production. A major portion of the world potassium production is consumed in making of fertiliser; however, other uses include those in pharmaceutical, glass, ceramic, food and detergent industries. The availability of soluble potash minerals (sylvite, kainite and carnallite) in different countries like Canada, USA, Israel and Russia makes them as major potash producers, whereas agricultural-based countries like Liberia, Somalia, Central African Republic, Thailand, Indonesia and Malaysia including India meet their potassium requirement through import only. On the contrary, the availability of huge potassium-bearing rocks/minerals (like nepheline syenite, feldspar, mica, glauconitic sandstone) in these countries containing around 4–17% K2O would be a prospective for commercial production of potassium. The potassium recovery from minerals/rocks is very complicated due to the uniform distribution of potassium throughout the crystal structure. Different physico-chemical separation methods like bioleaching, chemical leaching, flotation and roast leaching have been discussed for the successful recovery of potash values from these rocks/minerals. However, the recovery of potassium from hugely available seawater using the chemical precipitation, solvent extraction, membrane separation and ionic exchange methods is not cost-effective due to the low concentration of potassium.
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Silicate rock Verdete, collected in the central region of Minas Gerais state (Brazil) and composed mostly of micas (glauconite and muscovite) and tectosilicates (K-feldspar and quartz), was hydrothermally treated with several reactants in order to release and recover potassium. The hydrothermal products were characterized by flame photometry, XRD, XRF, SEM and EDS. Treatment with sulfuric acid was effective to break the crystal lattice of micas before 1 h of reaction and recovered 24% of potassium in the form of sulfates. The K-feldspar appears to have remained intact during the process. Treatment with a Ca(OH)2 (86 wt.%) - CaCO3 (14 wt.%) mixture did not consume the micas, but K-feldspar was gradually consumed over the 24 h reaction period. The K recovery was probably due to a concurrent hydrolytic framework dissolution of K-feldspar mediated by OH− ions and by the exchange of K+ with Ca2+. The K-bearing species are carbonaceous materials with variable K+/Ca2+ ratios, such as K2Ca(CO3)2.
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Various aspects of the reaction of phosphoric acid with muscovite mica at 250, 300, and 350-degrees-C were studied with a view to understanding the nature of such reactions, particularly (i) the reaction kinetics, (ii) the relation between muscovite dissolution and polymerization of phosphoric acid, (iii) the probable mechanism of reaction, and (iv) the nature of the residue. Solubilization of the K+ ion from muscovite was observed to be linearly dependent on the degree of dehydration of the system as well as the average chain length of the poly(phosphoric acid) formed. It is suggested that the breakdown of the complex muscovite structure is due to attack by OH- ions, which are produced when phosphoric acid polymerizes; oxide bonds are cleaved forming M-OH and M-O-P bonds, and the elimination of water from other P-OH groups results in polyphosphates. The reaction product consists of soluble and insoluble amorphous polyphosphates that form a coating over the core of unreacted mineral.
Until recent years knowledge of chemical processing was descriptive and qualitative. In 1810 modern chemical theory was born and process description became quantitative. Then about 1900 the quantitative engineering approach was developed, first for physical changes, called the Unit Operations, and somewhat later for chemical operations. This we call the American approach. In 1957 European chemical engineers brought together the design of chemical and their related physical operations under the name of Chemical Reaction Engineering, or CRE. This approach and name received practically universal acceptance. Today the methods of CRE are widely used in the processing of biochemical and all sorts of other systems, This talk wanders through this development.