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A case study of flotation of chalcopyrite-molybdenite ore using a mixture
of pyrogallic acid and o-isopropyl ethylthiocarbamate: Insights from
ab initio calculations
Yuanjia Luo
a
, Leming Ou
a
, Wei Sun
a
, Haisheng Han
a
, Jianhua Chen
b
, Jian Peng
a,
⇑
a
School of Minerals Processing and Bioengineering, Central South University, Changsha 410083, China
b
School of Resources, Environment and Materials, Guangxi University, Nanning 530004, China
article info
Article history:
Received 10 February 2023
Received in revised form 31 March 2023
Accepted 5 April 2023
Keywords:
Flotation
Chalcopyrite-molybdenite
Novel process
Sodium sulfide
Desorb
abstract
To date, lime-free flotation processes are increasingly advocated due to many problems caused by the
traditional process using lime (CaO) as a depressant and xanthate (BX) as a collector. In this work, we
developed a combination of pyrogallic acid (PGA) and o-isopropyl ethylthiocarbamate (Z200) as an
eco-friendly alternative to the lime-xanthate process for chalcopyrite-molybdenite (ccp-mlb) flotation.
Roughing flotation experiments showed that this process yields ccp-mlb concentrate with similar recov-
ery and grade of Cu and Mo as in the traditional lime-xanthate process. However, the huge advantage of
the novel process is that it requires ten times lower dosages of sodium sulfide in the subsequent separa-
tion of ccp-mlb concentrate. The ab initio calculations explained this trend at the atomistic level, where
hydrogen sulfide ions could desorb collector more easily from the Z200@ccp complex compared with
BX@ccp. The same method could be applied for environmentally friendly separations of other resources
with similar composition.
Ó2023 The Society of Powder Technology Japan. Published by Elsevier B.V. and The Society of Powder
Technology Japan All rights reserved.
1. Introduction
Porphyritic copper-molybdenum ores are the main mining
resource for chalcopyrite and pyromorphite, which account for
about 50% of global production of pyromorphite and 75% of global
production of copper [1,2]. Due to the excessive use of lime, which
has led to issues like pipeline jamming, environmental disruption,
poor selectivity, foam viscosity, difficult recovery of precious met-
als, and treatment with pyrite in dressing plants [3–5], the tradi-
tional process using lime (CaO) as a depressant and xanthate
(BX) as a collector is becoming increasingly limited. As a result,
lime-free flotation processes are gaining more and more attention.
Therefore, a new reagent scheme is urgently needed to be found to
separate the chalcopyrite and pyromorphite.
The beneficiation of copper-molybdenum ores is generally car-
ried out by a two-stage flotation method. The first stage uses thiol
collectors to recover copper sulfide and pyromorphite as a mixed
concentrate, and the second stage uses collectors to selectively
recover copper sulfide and pyromorphite [6,7]. Generally, in the
second stage but due to the similar floatability of pyromorphite
and chalcopyrite, their separation efficiency is low [8,9]. Over the
decades, researchers have developed a variety of inorganic and
organic reagents to inhibit copper sulfide in the flotation of
pyromorphite. Examples include lignosulphonates [10], disodium
carboxymethyltrithiocarbonate [11], pseudo-glycolythiourea acid
[12], L-cysteine [13], thiocarbonohydrazide [14], poly
(acrylamide-allylthiourea) (PAM-ATU) [15], O-carboxymethyl chi-
tosan [16], rhodanine-3-acetic acid [17], etc. However, these
agents have been studied almost exclusively in laboratory experi-
ments and have not been widely used on an industrial scale, and
the NaHS and Na
2
S are still the commonly used copper inhibitors
[18].
In this paper, we describe the flotation separation of pyrite-
containing porphyry Cu-Mo sulfide ore from Dexing, Jiangxi pro-
vince. In the first stage, we selected a suitable, environmentally
friendly reagent to depress the floatability of pyrite and provide
maximum recovery of chalcopyrite and Mo-rich minerals. In the
second stage, our focus was on improving the efficiency of sodium
sulfide (Na
2
S) to remove the collector from the chalcopyrite surface
and, eventually, to reduce Na
2
S dosage. Herein, we developed a
new process for the instant, environmentally friendly flotation of
chalcopyrite/molybdenite ore that combines PGA, Z200, and
Na
2
S. Compared with the traditional process that utilizes lime
https://doi.org/10.1016/j.apt.2023.104039
0921-8831/Ó2023 The Society of Powder Technology Japan. Published by Elsevier B.V. and The Society of Powder Technology Japan All rights reserved.
⇑
Corresponding author.
E-mail address: peng.jian@csu.edu.cn (J. Peng).
Advanced Powder Technology 34 (2023) 104039
Contents lists available at ScienceDirect
Advanced Powder Technology
journal homepage: www.elsevier.com/locate/apt
(CaO) as the inhibitor and butyl xanthate (BX) as the collector, the
proposed treatment was more efficient and required 70% less
Na
2
S to desorb the collector from copper concentrates. The key
intermolecular interactions responsible for lower consumption of
Na
2
S in instant flotation were revealed by ab initio calculations.
The combination of reagents proposed in this work is characterized
by low consumption of chemicals, low cost, and reduced environ-
mental pollution, and represents a promising separation strategy
that can be applied to similar mineral resources in the future.
2. Materials and methods
2.1. Minerals and reagents
2.1.1. Minerals
The studied pyritic Cu-Mo sulfide ore came from Dexing, Jiangxi
Province, China. The samples were submitted to a sequence of
coarse, medium, and fine crushing (Fig. S1,Supplementary
material), and the particles smaller than 2 mm were collected for
flotation experiments and mineralogical analysis. Samples were
stored in sealed bags at –4°C under a controlled atmosphere to
prevent oxidation.
The Cu, Mo, and S content in samples before and after flotation
was determined by ICP-MS. Other elements were quantified by
XRF. Table 1 presents the chemical analysis of the ore reveals. As
presented in Table 1, the content of the four most abundant ele-
ments is 0.37% Cu, 0.026% Mo, 2.52% S, and 2.79% Fe. The amounts
of other elements are too low to be utilized comprehensively. Min-
eralogical phase analysis was carried out on the ore samples using
a polarizing microscope. The mineral composition of the examined
ore is listed in Table 2. As presented in Table 2, the useful minerals
in this ore are mainly chalcopyrite (CuFeS
2
) and molybdenite
(MoS
2
), and their relative mineral contents are 1.28% and 0.054%,
respectively. Other metal sulfide minerals include pyrite with a rel-
ative content of 2.14% and trace amounts of galena and sphalerite.
Gangue minerals account for about 95.66%, mainly quartz, feldspar,
muscovite, biotite, chlorite, etc. The optical microscope image of
the raw ore is shown in Fig. 1. There are three main inlaid forms
of chalcopyrite. First, chalcopyrite is packed in quartz geodes of
quartz veins, and the grain shape and size are limited by the geodes
(Fig. 1(a)). Second, chalcopyrite is filled and metasomatized along
the edge of the pyrite, showing an irregular granular distribution
(Fig. 1(b) and 1(c)). Third, molybdenite is filled along the edge or
gap of chalcopyrite, and the two were closely associated (Fig. 1
(d) and 1(e)). There are also three main embedded forms of molyb-
denite in the ore. First, molybdenite fills the gaps in the quartz
grains or in the quartz geodes (Fig. 1(f) and 1(g)). Second, molyb-
denite is disseminated in gangue, and molybdenite is sometimes
seen contained in pyrite crystals (Fig. 1(h)). Third, molybdenite fills
along with ore fractures and replaces chalcopyrite (Fig. 1(i)).
2.1.2. Reagents
Butyl xanthate (BX, 92%), o-isopropyl ethylthiocarbamate
(Z200, 93%) and kerosene (95%) used as collectors in flotation
experiments were purchased from Zhuzhou Chemical Plant, China.
Terpineol (95%) was obtained from Guangxi Long Chemical Co.,
Ltd., China and used as a foaming agent. The pH was adjusted by
CaO (p.a., Aladdin Reagent Co., Ltd., Shanghai, China) and Na
2
CO
3
(p.a., Tianjin Kemio Chemical Reagent Co., Ltd.). Pyrogallic acid
(95%) and Na
2
S (95%) used as inhibitors were purchased from
Shanghai Mailin Biochemical Co., Ltd., China. All solutions were
prepared on the day of the experiment. Tap water was used for
all tests.
2.2. Flotation experiments
The flotation experiments on a laboratory scale were done on
XFG-type flotation device (Jilin Exploration Machinery Plant,
China) equipped with a 3.0 L flotation cell. The rotor speed was
1500 rpm. For each flotation test, 1000 g of ore was mixed
with 2.0 L of tap water in a flotation cell and stirred for 3 min
to form a homogeneous slurry with a grind fineness of 0.074 mm
(70% particles). Flotation tests were performed at 25 °C. The con-
centrates were separated by filtration, dried and weighed, and then
analyzed for the content of Cu, Mo and S. The recoveries were cal-
culated from the distribution of elements in concentrate and tail-
ing and the grade of the product. The amount of Cu, Mo, and S
was used to represent the grade and recovery of ccp, mlb and
pyr, respectively. Each flotation experiment was performed in trip-
licates, and final values were reported as means where error bars
represent standard deviations.
2.3. Ab initio calculations
Ab initio quantum chemistry calculations were performed with
the CASTEP module in BIOVIA Materials Studio to compare the
adsorption affinity of hydrogen sulfide ion, Z200 and BX on ccp
(112) surface. The GGA algorithm with PW91 functional was
selected [19–22]. During the calculation, TS method for dispersion
correction were added [21–23], and ultrasoft pseudopotentials
were adopted [24–31]. To model chalcopyrite, the cutoff energies
(E
cut-off
) was set to 351 eV and the quality of k-points was set to
fine [19]. Medium quality was selected for geometry optimization
in CASTEP.
The structure of ccp crystal cell (Fig. 2(a)) was retrieved from
American Mineral Crystallographers Database (AMCSD) [32–34].
The (112) surface of ccp was split from the crystal cell since it
has low surface energy [35] and packed into 2 21 supercell
surrounded by 30 Å of vacuum. The optimized structure of con-
structed ccp (112) surface is shown in Fig. 2(b). The 3D structures
of Z200, HS
–
, and BX were pre-optimized at k-point of gamma in a
30 30 30 Å cell and then allowed to interact with ccp surface.
Their configurations in the adsorbed state are shown in
Table 1
The elemental composition of ores (mass %).
Element Cu Mo S Fe Pb Zn Cl Au(g/t) Rb Sr
Content 0.37 0.026 2.52 2.79 0.028 0.037 0.027 0.11 0.023 0.052
Element Ag(g/t) Ba Ce Na
2
O MgO Al
2
O
3
SiO
2
K
2
O CaO P
2
O
5
Content 0.04 0.048 0.054 4.49 1.97 15.48 62.61 5.02 3.96 0.2
Table 2
The mineral composition of raw ore sample (mass %).
Minerals Content Minerals Content
Chalcopyrite 1.28 Muscovite 5.43
Molybdenite 0.054 Biotite 5.18
Pyrite 2.14 Calcite 0.18
Sphalerite 0.009 Epidote 0.015
Galena 0.005 Chlorite 0.99
Quartz 27.37 Apatite 0.33
Feldspar 56.16
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
2
Fig. 3(a)-(c). The two surface ccp layers were optimized while the
other four bulk layers were restricted.
The adsorption energy (
D
E) was obtained by the following
formula:
D
E¼Eðccp surface þadsorbateÞEðccp surfaceÞEðadsorbateÞ
ð1Þ
The more negative
D
Evalues indicate stronger interaction
between the ccp surface and the adsorbate.
As chalcopyrite is a sulfide mineral and is relatively hydropho-
bic, the calculations were carried out under vacuum conditions.
3. Results and discussion
3.1. Roughing flotation of chalcopyrite-molybdenite ore
Flotation efficiency of chalcopyrite (ccp) and molybdenite (mlb)
was improved by grounding the samples to obtain fine-grained
particles (Fig. 4). Fig. 4(a) and 4(b) illustrate the traditional
chalcopyrite-molybdenite-pyrite (ccp-mlb-pyr) bulk flotation and
the new ccp-mlb instant flotation, respectively. Fig. 4(c) shows
the process flow applied in the second step of ccp/mlb flotation
separation test.
In this study, we applied these two approaches to recover chal-
copyrite (ccp) and molybdenite (mlb) from ccp-mlb ore. In a con-
ventional approach, ccp and mlb were recovered simultaneously
using sodium carbonate as pH regulator, CaO as the inhibitor and
butyl xanthate (BX) as the collector. The efficiency of the tradi-
tional approach was compared with the novel, instant flotation
process that utilizes pyrogallic acid (PGA) as the inhibitor and
o-isopropyl ethylthiocarbamate (Z200) as the collector. Both
approaches require the addition of sodium sulfide as the ccp flota-
tion inhibitor to separate ccp from mlb.
The results in Fig. 5(a) and Fig. 5(b) show the influence of CaO
and BX dosage on the grade of Cu and Mo in ccp-mlb bulk concen-
trates and S in pyrite (pyr) concentrate. The corresponding raw
Fig. 1. Optical microscopy and SEM images of the untreated ore.
Fig. 2. The crystal cell of chalcopyrite (a) and the optimized (112) surface of
chalcopyrite (b).
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
3
data are given in Table S1 and Table S2. An increase in CaO dosage
up to 2000 g/t improved the grade of Cu and Mo reaching 10.88%
and 0.7%, respectively. The grade of Cu was almost independent
of CaO dosage, while the grade of Mo decreased at dosages above
2000 g/t. This result indicates that the optimum CaO dosage for
Mo recovery from ccp-mlb ore is 2000 g/t.
Keeping the amount of lime fixed at 2000 g/t, we studied the
influence of BX dosage on the recovery and grade of ccp-mlb con-
centrate. At 60 g/t BX, the recoveries increased to 82.05% and
75.12%, respectively. Further increase in BX negligibly influenced
the recovery of the minerals from ore. Corresponding grades of
Cu and Mo were 11% and 0.7%, respectively. Moreover, the
recovery and grade of S in pyr concentrate were 67% and 25%.
In a new approach, we have used environmentally friendly PGA
and Z200 as replacements for CaO and BX. The results in Fig. 5(c)
and Fig. 5(d) show the influence of reagent dosage on the recovery
and Cu-Mo grade in ccp-mlb bulk concentrate and S in pyr concen-
trate. The raw data can be found in Table S3 and Table S4. The
recoveries of two transition metals in bulk concentrate increased
with the increase in PGA dosage up to 300 g/t. A higher amount
of PGA resulted in a slight recovery drop. Furthermore, the influ-
ence of Z200 dosage was studied by keeping the amount of PGA
at this optimum level. Fig. 5(d) shows that the recovery of Cu
and Mo increases up to 30 g/t Z200, and remains constant at higher
dosages. Under optimum reagent dosage (300 g/t PGA and 30 g/t
Z200), the recoveries of Cu and Mo in ccp-mlb bulk concentrate
reached 87% and 89% with grades of 12% and 0.8%. More-
over, the recovery and grade of S in pyr concentrate were 68%
and 33%.
Comparing the results of the first step of conventional (A) and
instant ccp-mlb bulk flotation (B), it is clear that the dosages of
inhibitor and collector are significantly lower for the instant flota-
tion process (2000 g/t and 60 g/t of CaO and BX vs. 300 g/t and 30 g/
t of PGA and Z200). The Cu and Mo recoveries in concentrate B
were higher by 5% and 3%. The Cu grade was also slightly
improved in concentrate B while Mo grade was similar in both con-
centrates. The recovery of pyr concentrate was increased by 1% in
an instant flotation, and the S grade was improved by 8%. We con-
clude that the combination of novel, eco-friendly flotation reagent
results in the efficient separation of ccp-mlb and pyr.
The addition of sodium sulfide in a bulk concentrate effectively
depresses chalcopyrite flotation and allows the separation of chal-
copyrite from molybdenite in ccp-mlb bulk concentrate. In this
study, we optimized the dosage of Na
2
S to maximize recovery
and grade of the minerals. The results of ccp-mlb bulk separation
were compared for traditional flotation (concentrate A, Fig. 6(a))
and a new, instant flotation scheme (concentrate B, Fig. 6 (b)).
The raw experimental data can be found in Table S5 and Table S6.
Fig. 3. The optimized geometries of Z200 (a), hydrogen sulfide ion (b), and BX (c).
Fig. 4. The process scheme illustrating (a) traditional ccp-mlb-pyr rough flotation; (b) newly proposed ccp-mlb instant flotation and (c) flotation separation of ccp and mlb.
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
4
The results in Fig. 6(a) show that the grade and recovery of both
minerals increased up to 3000 g/t Na
2
S, while higher concentra-
tions have negligible impact on the separation efficiency. The effec-
tive separation of concentrate was confirmed by high ccp recovery
(88.49%) with good Cu grade (10.99%), as well as with excellent
mlb recovery and Mo grade (89.12% and 5.74 %, respectively).
The trends in recovery and grade of both minerals from con-
centrate B (Fig. 6(b)) were consistent with Fig. 6(a). However, only
300 g/t Na
2
S was necessary to reach the same separation effi-
ciency as in concentrate A. The corresponding recovery and Cu
grade of ccp were 88.03% and 11.87%, while that of mlb and Mo
grade were 90.85% and 6.08%. These results indicate that the
selection of flotation inhibitor and collector applied in the first
step of ccp-mlb flotation have a significant influence on the
dosage of chalcopyrite depressant required for the efficient sepa-
ration of bulk concentrate in the second step. Instant bulk flota-
tion with eco-friendly PGA and Z200 requires only one-tenth of
Na
2
S dosage from that of the conventional lime-xanthate process
in the subsequent separation step.
3.2. Closed-circuit experiments
The type of the reagents and their dosages optimized in rough-
ing flotation experiments were taken as the starting point for
closed-circuit flotation tests. The experimental procedure for this
test is illustrated in Fig. 7. In the first step, Cu-Mo concentrate A
was obtained by instant flotation using the combination of
Z200 + PGA. In the cleaning stage, only PGA inhibitor was used.
Fig. 5. Grades and recoveries of Cu, Mo in ccp-mlb bulk concentrate and S in pyr concentrate as a function of CaO (a) and BX (b), PGA (c) and Z200 (d) dosages.
Fig. 6. The effect of Na
2
S dosage on the separation efficiency of ccp-mlb bulk concentrate obtained by (a) conventional lime-xanthate process and (b) instant PGA-Z200
flotation.
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
5
Afterward, asynchronous flotation according to a conventional
lime-xanthate scheme gave Cu-Mo concentrate B. Since the con-
ventional separation of Cu-Mo/S requires large amount of lime rel-
atively large loss of Mo is expected. As shown in 3.1, two kinds of
Cu-Mo mixed concentrates have a similar grade and recovery of Cu
and Mo but concentrate A requires a significantly lower amount of
sodium sulfide to desorb Cu from the collector. Using the flow sys-
tem illustrated in Fig. 7 it is possible to adjust the system to obtain
as many instant flotation Cu-Mo concentrates as possible, which is
beneficial to the recovery of Mo.
Under the closed-circuit experimental setup (Fig. 8) and using
the same scavenger and cleaner as in roughing separation of ccp
and mlb from ccp/mlb concentrate it was possible to halve the
amounts of reagents. Moreover, sodium sulfide dosage for Cu-Mo
concentrate A (Fig. 8(a)) was only one-tenth of the dosage for
Cu-Mo concentrate B (Fig. 8(b)). The final yields of three minerals
obtained by closed-circuit flotation are listed in Table 3. High
recoveries of Cu and Mo from corresponding concentrates
(89.81% and 80.50%) and excellent grades (23.40% and 45.50%,
respectively) have confirmed that the newly developed reagent
scheme efficiently separates ccp from mlb. This is provided by
closed-circuit experiments designed to exploit the fact that the
concentrates obtained by traditional and instantaneous flotation
require different amounts of sodium sulfide for collector desorp-
tion from copper concentrates.
3.3. Ab initio calculations
Ab initio quantum chemical calculations provided insights into
the adsorption of Z200 and BX on chalcopyrite (112) surface and
the affinity of hydrogen sulfide ions (HS
–
) to desorb collectors from
the surface. Table 4 lists the adsorption energies (
D
E) between
Z200 and BX and ccp surface, as well as
D
Ebetween four HS
–
and the adsorption complexes of ccp with Z200 and BX. The lower
D
Efor the adsorption of HS
–
ions on the surface of Z200@ccp
Fig. 7. Experimental procedure for closed-circuit Cu-Mo flotation.
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
6
complex confirms a higher affinity of sodium sulfide to desorb
Z200 and lower dosages of reagent needed for separation of ccp
and mlb from concentrate.
The optimized configurations for the adsorption of hydrogen
sulfide ions to Z200@ccp and BX@ccp complexes are shown in
Fig. 9 (a) and Fig. 9 (b). The binding of HS
–
ions to Z200@ccp cre-
ated a Cu-S bond length of 4.64 Å. Accordingly, the lengths of
Cu-S bonds of BX@ccp became 4.76 Å and 3.90 Å. These Cu-S dis-
tances are far larger than the covalent radius of Cu-S, thus Cu-S
bonds all have broken down. This result suggests that both collec-
tors could be efficiently desorbed by hydrogen sulfide ions formed
as a product of Na
2
S hydrolysis.
Fig. 8. Experimental procedure for closed-circuit Cu/Mo separation.
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
7
Fig. 10 (a) and Fig. 10 (b) show the electron density differences
for Z200@ccp and BX@ccp complexes after the adsorption of four
HS
–
ions. The regions with blue color are parts of the adsorption
complex with the increased electron density compared with
pristine adsorbent, while red-colored regions are the ones with
the decreased electron density. As presented in Fig. 10 (a) and
Fig. 10 (b), this electron transfers in the direction of Cu-S were
not observed in case of HS
–
adsorption to Z200@ccp and BX@ccp
complexes, which indicates that both collectors were successfully
desorbed from ccp (112) surface.
In summary, ab initio calculations explained the relative affinity
of HS
–
ions to desorb Z200 and BX collectors from ccp on an atomic
level. This result provides insights into the molecular mechanisms
of instant flotation and explains why lower quantities of Na
2
S are
required to desorb collector from Z200@ccp complex.
4. Conclusions
In this paper, we applied an eco-friendly combination of pyro-
gallic acid (PGA) and o-isopropyl ethylthiocarbamate (Z200) to
replace the traditionally used lime-xanthate process. To efficiently
separate ccp from mlb, we used sodium sulfide as the inhibitor.
Roughing and closed-circuit experimental setups were applied to
separate pyrite-rich CuS/MoS ore. Compared with ccp-mlb concen-
trate obtained by conventional process, concentrate obtained by
instant PGA-Z200 flotation required significantly lower dosages
Table 3
The results of closed-circuit flotation experiments (given in mass %).
Product Yield Grade Recovery
Cu Mo S Cu Mo S
Chalcopyrite concentrate 1.42 23.40 0.25 23.98 89.81 13.65 13.51
Molybdenite concentrate 0.046 0.78 45.50 32.75 0.097 80.50 0.60
Pyrite concentrate 4.31 0.38 0.0064 46.75 4.43 1.06 79.96
Tailings 94.224 0.022 0.0013 0.16 5.67 4.79 5.93
Raw ore 100.00 0.37 0.026 2.52 100.00 100.00 100.00
Table 4
Interaction energies between Z200 and BX collectors and ccp (11 2) surface, and the
interaction energies between HS
–
ions and Z200@ccp and BX@ccp complexes.
Mineral surfaces (Adsorbent) Adsorbate
D
E(KJ/mol)
ccp Z200 56.60
BX 157.00
Z200@ccp 4 HS
–
446.65
BX@ccp 4 HS
–
310.81
Fig. 9. The optimized configuration of Z200@ccp (a) and BX@ccp (b) complexes upon binding of four hydrogen sulfide ions.
Fig. 10. The difference in electron densities after four HS
–
ions on Z200@ccp complex (a) and four HS
–
ions on BX@ccp complex (b).
Y. Luo, L. Ou, W. Sun et al. Advanced Powder Technology 34 (2023) 104039
8
of Na
2
S to reach the same separation efficiency. The closed-circuit
setup was designed to exploit this difference, which resulted in
excellent recoveries and grades of ccp and mlb (89.81% with
23.40% Cu, and 80.50% with 45.50% Mo). Ab initio calculations pro-
vided insights into the intermolecular interactions responsible for
the key features of the novel flotation process. The results revealed
that HS
–
ions have a higher affinity to desorb collector from
Z200@ccp complex compared with BX@ccp complex. This process
scheme has potential for future separations of other ores with sim-
ilar mineralogical properties.
Declaration of Competing Interest
The authors declare that they have no known competing finan-
cial interests or personal relationships that could have appeared
to influence the work reported in this paper.
Acknowledgments
This work was supported by the financial support from the
National Key R&D Program of China (No. 2022YFC2904502,
No. 2022YFC2904501), Hunan High-tech Industry Technology
Innovation Leading plan (No. 2022GK4056); This work was
supported in part by the High Performance Computing Center of
Central South University.
Appendix A. Supplementary material
Supplementary data to this article can be found online at
https://doi.org/10.1016/j.apt.2023.104039.
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