PH12018050174A1

PH12018050174A1 - Enhanced methods of extracting precious metals and methods of testing

Info

Publication number: PH12018050174A1
Application number: PH12018050174A
Authority: PH
Inventors: Herman D Mendoza
Original assignee: Univ Of The Philippines Diliman
Priority date: 2018-04-18
Filing date: 2018-04-18
Publication date: 2019-11-11
Also published as: WO2019203667A1

Abstract

This invention provides the extraction of precious metals using enhanced gravity concentration-flotation processes of extracting the metals from the ore. Moreover, wastewater from the extraction stage is treated before being discharged to the environment using a combination of zeolite and cocopeat, materials that are readily available to small-scale miners. Methods of testing the foregoing are likewise provided.

Description

The gold precipitates GP is mixed with borax and undergo refining using a blow i, torch to produce the final product — a gold bead.

The tailings from final tails thickener 28 is discharged into final tailings settling pond 51. The solids are allowed to settle producing clear water CW that can be - disposed to a natural body of water like river R. The solids from final tailings : settling pond 51 and the barren solids BS from filter press 32 and pressure filter = 45 can be packed into sacks 56 or similar containers for final disposal or may be ow utilized as additional support structures to counter erosion or even for © landscaping. =

The washings and barren solution (considered as the wastewater from pressure filter 45) are placed in neutralization tank 52 to neutralize the pH using sodium hydroxide before discharging to the wastewater treatment compartment 53.

The neutralized barren solution is then discharged from wastewater treatment's 53 first compartment 531. The solution then flows through cocopeat — zeolite layer 54. The treated water at the bottom of cocopeat-zeolite layer 54 goes to a holding area/empty compartment 55. When this compartment becomes full, it overflows to second compartment 532 with a cocopeat-zeolite layer for the second stage of treatment. The treated water from the second stage proceeds to third compartment 533 with 100% zeolite for the final stage of treatment. The treated water from the third stage, together with clear water CW, is then discharged to river R.

ee ——— 8

Gravity Concentration Tests hs

The recovery of free gold was conducted through the gravity concentration : provided by this invention. Gravity concentration tests were done using an L40 -

Falcon gravity concentrator. A 32-factorial experiment design was performed to = determine the operating parameters, particularly the size of the feed (in terms of wo the percent passing Pso) and the gravitational force of the concentrator (G Force). =

Both parameters were set at three levels as shown in Table 1. -

Table 1. Design of experiment for the Falcon gravity concentrator - parameter (Goce | 0 | 00 | 10

For each run, 10 kilograms of ore was used as feed and was set at 10% solids.

The feed to the concentrator has an average head assay of 7.05 gpt of gold. The operating water fluidization pressure for the concentrator was also maintained at 6 psi. The results are presented in Table 2.

Table 2. Results of the gravity concentration experiments.

Ps, % mass Conc. Grade, Concentration R 0 microns pull-out gpt Au Ratio ecovery,

CT M6 | 0 | os | AZ| 4 | Wo 2 im | fe | os | aor | teas | 899 3] | Ww | om | pd | 2d | 60 4] 120 | 100 | 08 | 5498 [| 12048 | 640

[5] 6 | ew | oo | aww | few | w2 6] 7 | dw | os | me | feds | 18 7] a | 0 | i | 6% | 7246 | 13

8] dw | dw | on | dem T3889 - 9] 5 | we | os | mim 196.08 3

The results in Table 2 show variable yield for the different settings of Pg and G =

Force. As can be seen, Run 4 resulted the highest recovery of gold at 64% with - considerable concentrate grade of 543 gpt of gold. -

A graph of recovery versus G Force is illustrated in Figure 2. A general trend can o be observed wherein finer particle sizes require lower G Forces to attain higher oy recoveries. o

Recovery vs. G Force 90 pn i ore en | - A= 75microns 80 in ns SU : - Q Ce _..| — » — 106 microns 70 Ce i oe 120 microns § 60 w Jol _ J I Cg oo 9 §Q Fo N ¢ es - ~>é& oe : . i <4 @ . K ~< ; : 40 bone Ne ie LN So _ es ed 30 FN cee SK I . . i Ce Ne FE “ PR

CY & To TX ~ ~, J -A 50 70 9G 110 130 15C 170

G Force

Figure 2. Recovery versus G Force.

Flotation Tests

A Taguchi design of experiment (DOE) was performed to determine the parameters for flotation. The parameters considered were feed size (Pso), pH level, frother type, collector type and dosage. Each parameter was set at three ~ levels or types as shown in the Table 3. -

Table 3. The Taguchi design of experiment set for flotation tests -

FeedPw | 120um | 106um | 75um | ~ -

For the collector and frother machines used, CMS is an Australian brand of thionocarbamate collector, SIBX is a locally produced xanthate and S701 is a

Nasaco dithiophosphate collector. Nasfroth 240 is a glycol based frother,

Nasfroth 626 is an alcohol based frother both coming from the Nasaco brand, and the IF6500 is an Australian brand of glycol-based frother.

For each run, a sub-sample was ground to the target flotation feed size. Flotation was conducted in a 5L flotation cell at 40% solids. The concentrates were recovered in 4 batches called R1, R2, R3 and R4. A staged flotation was employed to have an idea of the laboratory-scale flotation time appropriate for the ore.

A head sample is collected before dosing the pulp. Conditioning time was set to 2 minutes per stage to allow the reagents to mix with the pulp. In R1, concentrates were scraped from the froth phase into a sample pan for 1.5 minutes. The aeration is then turned off, another dose of the reagent is added

1 mn, and a new pan is prepared for the next stage before turning the aeration on =O again. R2 to R4 followed a scraping time of 2 minutes, also following all the steps - mentioned. A tailings sample is also taken before the flotation vessel is = replenished with a new sample. The flotation testing conducted a total of 27 runs. s The experiment highlighted the significant factors that affected the gold flotation - concentrate grade and recovery of the ore sample namely, collector dosage and w pH. A summary of the flotation runs that yielded significant results is given in -

Table 4. ©

Table 4. Flotation run results = [mem Tome [5 Ln [Fr To [S27 os | 7s | 6 | 9 | Nao | SBX | 15 _ ois | fem | 06 | 9 | Nex | Sex | 15 0 | wis | mm | 8 | Neem | SEX | 10 [ws | em | i | 85 | Few | oWs | 15 ois | fees | 0 | 85 | NFoAb | CWS | 15 ee | we | 7 | 8 | Few | soi | 15 ois | sia | 5 | 6 | Few | ows | 10 [ws | wm | 75 | 9 | Neaw | ows | fo ore | mss | 7% | 9 | Neem | ows | 10

Based on the results, the best pH for operation was at pH 9. Frother types were not very significant to the yield but most of the high recovery runs used the alcohol based NF626. Collector type also did not show any significance, showing both thionocarbamate CMS and xanthate SIBX fit for the process. Relating this to the appropriate collector dosage, 10gpt is the better dose for runs that used CMS while 10gpt to 15gpt was appropriate for runs with SIBX. Feed size showed the least significance as all particle sizes considered yielded similar concentrate grade and recovery. A better visualization of the relationship between the = different factors is illustrated in Figure 3. fa @) (®) ht

Feed peo vs Recovery Feed pso vs Grade TE 3100 or e 84.00 fom y.,. 81.00 m I. re ou. bin ne nl

Ee a Ea 75.00 Lo i ok ev fee a] pe | - 75 105 120 75 106 120 oy me -

GRIND, Ro GRIND, MICRONS La (c) (d) EE vsRecovery

Collector Type Collector Type vs Grade 50.00 120.00

La J Fas) —

L. 8408 | Soo :

Bo ed a

Bean a

Pati oe ae on Lela] » 4000

CVI S73. $103 wr (e) to vs Recovery

Frother Type Frother Type vs Grade 85.00 ages 140180 rT z 116.23 84.00 ToT a 24 83.00 .... 60.00 Gp 82) 81.00 120 40.00

‘ 13 = (9) (h) re

BOLD vv mom sisi toes ms 0 tt i” vr, w- 100.00 tT To 75.00 + EES - £ 80.00 : i po # sooo i wn vod - = 0.00 - ee oo — £ 45.00 be SR nnn ri g TTT ETT - - = goo I . g om i : on 15.00 § woo . a 000 . os , 5 LY : Lm os erm , i BN

PH LEVEL © PH LEVEL ~

H on a 1000 ,, 12000 . : - \ = 20.00 . - EAEN ¢ rs E 000 $m | Lo... £ j CEH oy go z = = 00 go 1 0.00 : Nn 8 1 | SF 0 15 5 10 1s

COLLECTOR DOSAGE, GRAMS PER TONNE COLLECTOR DOSAGE, GRAMS PER TONNE

Figure 3. Relationships of the different factors.

Flotation time was determined to be between the 5.5 to 7.5 minutes period. A 5 graph showing the cumulative recoveries per run is shown in Figure 4. The highest recovery was obtained after 5.5 minutes and only a small increase is observed when extended to 7.5 minutes. This data can be used as the baseline residence time for the scale up in the pilot-scale.

FLOTATION RECOVERY CURVE

100 5 20 8 : Go EE ———— i 2 a0 SE g = 20 S—— a — 0 frre 0 1 2 3 4 5 6 7 8 flotation Time, minutes

Figure 4. Recovery curve from the batch flotation test. =

Extractive Metallurgy Tests =

The recovery of gold from flotation concentrate and purification of gold from the r gravity concentrator were conducted through leaching and precipitation. An - s extraction process was developed that fully deviates from the use of cyanide and mercury. Several laboratory tests have been conducted to establish the ie extraction procedure and operating parameters. The laboratory tests resulted in = about 80% recovery of gold. - a. Lixiviant Composition 0

Preliminary experiments were conducted to determine the type of lixiviant to be used for leaching. The types of lixiviants compared were thiourea, thiosulfate, iodine, and chlorine (see Figure 5.). The results showed that iodine yielded the highest recovery. However, iodine is expensive therefore is uneconomical to use.

Chlorine resulted the second highest recovery. As it is economical to use, chlorine was considered and developed for gold leaching (via chlorination process).

Comparison 90 80 A —————s sto ——

S50 Sf won Chiorinat ion dw —a—cyanidation

WL i iodee 2 Va main Thiosulphate ae 0 50 100 150 200

Time, minutes

Figure 5. Preliminary experiments comparing various types of lixiviants. ~

Prior art that utilized the chlorination process used the following reagents: x sodium hypochlorite, calcium hypochlorite, sodium chloride and hydrochloric © acid. The recipe for chlorination suitable for sample ore was then determined i (shown in Table 5). The recovery of gold from the varying recipes was plotted - and illustrated in Figure 6. N

Table 5. Types and dosages of lixiviants for leaching -

Designation Vola of o

Wivint | F0_| Wael | NaOCl | AGI | CaoCh [A [fom | 18 | 9 | & | 6 | - [8B | fom | | 8 | s% | 1H | 3

Cc [dem | | fo | sm | 7

D0 [eel [ - [3 | we | & | W_]

The highest recovery, about 60%, was achieved using recipe B. This is due to the differing properties of the complexing ligands such that the calcium hypochlorite is more stable than sodium hypochlorite and has a higher chlorine content while sodium hypochlorite has kinetics twice that of calcium hypochlorite. Based on this, the succeeding leaching experiments used the following recipe: 18% H-0, 8% NaCl, 58% NaOCl, 16% HCI, and 3% Ca(OCl)..

Different Lixiviant Recipes ~ =A =f === =D - 70 . 50 / ; iT

LP IN NC - $2177 ~ 0 50 100 150 a

Time, minutes

Figure 6. Effect of the various lixiviant recipes on the recovery of gold. =

A further test involving decreased amount of hydrochloric acid was conducted. It was found that decreasing the hydrochloric content up to 0.32% of the original concentration can still yield comparable results as compared to the concentrated mix. Hydrochloric acid is added to oxidize the hypochlorite ions from the complexing ligands (sodium hypochlorite and calcium hypochlorite). The sample concentrate is refractory (wherein the gold is enclosed inside a sulfide mineral usually pyrite). Since the mechanism of dissolution of gold in chlorination is oxidation of the gold, the sulfide minerals would also be oxidized. When a sulfide mineral such as pyrite is oxidized, the sulfur in the mineral would be converted to sulfuric acid and can serve as an oxidizing agent to the hypochlorite ions, thus reducing the required amount of hydrochloric acid needed for hypochlorite oxidation. The decrease in the amount of hydrochloric acid was based on the resulting pH of the water-plus-hydrochloric acid mixture at the start of the leaching process. Figure 7 shows the result of HCI dilution on the recovery of gold.

Dilute HCI - = 1:1 HCEH20 WpH2 mpH4 - 900 = 80.0 i. 70.0

E800 {—— - $ s00 — - 300 ri 100 - So 00 4— | @ 1 oo

Time, 6th hour

Figure 7. Effect of HCI dilution on the recovery of gold. b. Mode of Addition of Lixiviant

Gold dissolution occurs in two stages. [Au+] is an intermediate species in the leaching reaction. The [Au+] that was initially formed is the species would react with the Cl- ions to form the more stable gold chloride [AuCl4] complex. However, if the concentration of [Au] increases rapidly, based on Le’ Chatelier's principle, the backward reaction would take place. This would slow down the rate of dissolution of gold and would lessen the recovery of the process. To ensure the continuous

Benguet Analysis 1 mA ng ne 0 BE nF G =H 120 100 § 80 60 § « 0 1 6th hour(final hour)

. oxidation of gold from the solids, the [Au*] concentration should be kept at a - minimum. This can be done by ensuring that the [Au+] formed would immediately ro react with CI- ions in the solution. The intermittent addition of Ca(OCI)2 and NaOC! > ensures that a steady supply of Cl- are available to react with the Au ions that are @ s formed throughout the leaching process. Bulk addition of the hypochlorite species tw might cause the amount of Cl- ions available to drop during the latter stages of the = process due to complexation of other metals to the chlorine ions. The design of ~ experiment and the results are summarized in Table 6 and Figure 8. Bulk addition + is adding all the leaching reagents at the start and waiting for a fixed amount of o time for the reaction to finish. The iGoli method is an incremental mode of addition. -

It involves adding HCI at the start, and then NaOCl is added incrementally until the - end of leaching time.

Table 6. Results of the mode of addition, presence of NaCl, and pre-treatment on

CA Tek Tween wed 0]

C | Buk | wih | umwoased | 65 0 | ei | wh | wowed | 8

EE | Buk | wow | roased | 6

Fgh | wow | oesed | 9% sp A 1 el win | resed

Benguet Analysis 1 - ma na "oe aD RE BF LI] oH : 120 fe. 100 Ey i oe 60 20 toa 0 . 1 rz 6th h ous{final hour) Ao

Figure 8. Effect of the mode of addition of lixiviant on the recovery of gold. o 5c. Pulp Density

Tests were done to determine the pulp density that should be used in leaching.

The result illustrated in Figure 9 show that higher leaching recoveries were observed at lower pulp densities. This could be due to better kinetics since it would be easier to agitate pulps that have a lower pulp density. Better agitation of the pulp would make it easier for the leaching agents to diffuse into the surface of the gold particles.

Lowering the pulp density, however, will have a detrimental effect to the subsequent precipitation process. The dilution would decrease the effectivity of the precipitation process as shown in Figure 7. To achieve an optimum leaching recovery and its subsequent precipitation, the pulp density should be at 1.1421 g/mL.

Average Recovery - 8 38.7 % solids » 25% solids 18% solids Pd 800 re re i 708 e

Er — E doo] os 00 4] 00 |] os 1 Li

Figure 9. Effect of pulp density in the recovery of gold after 6 hours of leaching. - d. Roasting as Pre-treatment and the Effect of Hypochlorite Species =

Roasting was done as a pretreatment so that the sulfides in the ore sample = would be oxidized, thereby exposing gold locked in sulfide minerals. Based on ed tests conducted (results are shown in Table 7), roasting was found to be a significant factor to increase gold recovery. The tests also served as verification on the significance of sodium chloride in the process. It was found that sodium chloride does not significantly affect the recovery thus, can be omitted to decrease reagent cost.

Cost |_| 1819 31% | 000

Prelreaiment 0.006 0.166

Ne | os | 0% | 281 | 022 | 08%

Pretreatment*Lixiviant 1190 | 595 | 251 | 237 | 0045

Preeaiment NaC 0551

CivianNacl 2m | 15 | 251 | 05 | oan

Prereaiment Lan NaC

From Table 8, it was observed that the mixture of lixiviant contributed to the = increase in recovery. The effect of using different forms of hypochlorite, particularly sodium hypochlorite (NaOCl) and calcium hypochlorite (CaOClz) on = leaching recovery utilized the advantages of each form. Sodium hypochlorite has ~ a higher solubility and the greater availability of chlorine in calcium hypochlorite. -

These characteristics were major considerations in determining the effective pulp Fe density for leaching while keeping dilution at the minimum. Based on the results, = the NaOCI + Ca(OCl), mix resulted in the highest recovery. Therefore, the ~ lixiviant mix used in the succeeding leaching tests comprised of 48% NaOCl and = 2% Ca(OCl); of the lixiviant volume. =

Table 8. Effect of pretreatment, hypochlorite form and presence of sodium chloride

A | Uwoased | NaC | wow | 75

B | Rosed | NaC | wihow | @

C | Unomsed | WaOC+CalOC: | wiowl | 80

Tests were conducted to determine the optimum leaching time. The results are shown in Figure 10. It can be observed in Figure 10 that higher dissolution of gold occurred after 3 — 4 hours of leaching. This is because gold is oxidized before the competing sulphidic minerals. However, prolonging leaching beyond 4 hours decreased the dissolution. This suggests that gold may have been re-precipitated back to the pulp. Leaching must be conducted on the conditions that gold is being dissolved and must be stopped at the onset of the oxidation of other competing h minerals. Based on the observed trend, leaching time was preferred to be at 4 " hours. -

Leaching Recoveries per Hour = 50 Su 30 Pah meee QUI 1 Le ot NC @ 0 50 100 150 200 250 po

Hme, minutes LF

Figure 10. Effect of leaching time on the recovery of gold. f. Simultaneous Oxidation and Leaching

Due to SO2 emissions from roasting, an alternative pretreatment process was examined. Several oxidizing agents were considered. Acid washing using hydrochloric acid (HCI) was explored. However, as determined through Bruce method, no gold is locked within the carbonates of the ore. HCl is only effective in oxidizing carbonates and has no effect on sulphidic minerals. Sodium hydroxide (NaOH) or caustic soda was then studied. Previous studies showed that pretreatment of ore with pyrite lattice using caustic soda improved gold extraction as shown in Table 9.

Table 9. The effect of sodium hydroxide dosage and pretreatment time on = recovery 2 oe

A simultaneous oxidation and leaching method was established which is derived - from previous studies. The method is a combination of alkaline caustic soda > oxidation and the Igoli method. It involves a two-part process: 1) oxidation of » sulfides at alkaline conditions to expose the locked gold (as well as leach any free gold in the concentrate); and 2) leaching of the exposed gold. In this method, an amount of NaOCI, NaCl, NaOH, and Ca(OClI); is first mixed until homogeneity.

The Ca(OCl), was added to ensure that the strength of the hypochlorite is maintained even with the dilution of the concentrate slurry. The diluted concentrate was then added and the whole mixture was agitated for 2 hours. This is the oxidation part. After 2 hours, the pH was adjusted to a lower value (pH 8-9) using 6 M HCI. Upon reaching the desired pH, the leaching process was started by adding a fixed amount of Ca(OClI),, every 10 minutes for 4 hours. After 4 hours, the whole mixture was filtered. The filtrate (pregnant leached solution) was collected for precipitation. Aliquots were obtained every 30-minute interval, and were analysed for gold content. The result is illustrated in Figure 11. It can be seen in Figure 11 that the amount of gold leached increases until the 120th minute (2 hour) but decreases as leaching time is further increased. This may be due to the gold complex becoming unstable after the 120 minute. The

24 me, reagent NaCl increases the stability of the complex and the amount of NaCl at =O the 120% minute may not have been sufficient to keep the complex stable. =

The retrofitted operating parameters in the leaching process based on = simultaneous oxidation and leaching method tests are shown in Table 10. i.

Recovery of gold at various leaching times n

J Li ~~ TT 90 NT TT 70 bade Xe een fa

J i : i i ha

S 50 fone foinn eee eefe oe ere

C20 beeen md md err minnie xX i i : : : 0 i i i 1 NE —— i 0 50 100 150 200 250 300 350 400

Time, minutes

Figure 11. Recovery of gold using the simulatneousoxidation and leaching method at various leaching times.

Table 10. Modified Operating Parameters for Pre-treatment and Leaching [| Reagent [| %Volume | Time

NaOCl, 6% 60.62%

NaOH, technical grade 0.71%

Pretreatment. |e OCI (oxidation) 043% 2h 345% 585%

Ca(OCl): 5.72% resenng 345% 6 MHC 4.55%

Note that NaCl and Ca(OCl), are added incrementally to maintain the strength o of HOCI species in the solution as well as the stability of the gold auro complex > formed. Incremental addition was done every ten minutes until the 4 hour. = g. Precipitation ~ s Stage precipitation was performed using sodium metabisulfite (SMB S) and - ascorbic acid (HAsc). Initial tests were already conducted to determine the . feasibility of this process for gold recovery. These tests were successful in precipitating gold from pregnant solution after leaching. Using a blowtorch, = precipitates obtained were converted into gold beads thereby verifying the = viability of the process. With this, further optimization tests were conducted to - refine precipitation process. The parameters for the test are different concentrations for SMBS and HAsc, agitation, and pH. The stripped solution was then read using XRF and the results are shown in Table 11.

Table 11. Gold concentration of the stripped solution after the precipitation process

SMBS conc SMBS conc [Fgh | 5] Oo 19] 68] Hg [02] 43] 50] 35] [Tow | 85 42] ie] f67] low | 405] 84] a2] 73

A lower reading of the XRF is favored because it implies more gold was precipitated and only a small amount remained in the stripped solution. As expected, higher values of both the precipitating reagents (SMBS and HAsc) yielded a higher percent recovery. As for the pH, theoretically, a more neutral pH is favored for the ascorbic acid (HAsc) since within this pH range, it is in its = deprotonated form. The deprotonated form of HAsc is a stronger reductant as - compared to its other species. 2

The data in Table 11 were tested for analysis of variance (ANOVA) at full factorial to determine the parameters which has significant effect on gold recovery. The result is shown in Table 12, where factor A is the SMBS concentration, B is HAsc N concentration, C is the pH and D is agitation. =

The highlighted factors, namely HAsc concentration, agitation and the interaction = of both, are those with p-values less than or equal to 0.05 and were considered - significant. The pH did not seem to have significant effect on the precipitation process (ie. whatever the pH of the solution will be, the recovery of the process is the same).

Confirmatory tests were made based on half-factorial design. Unlike the previous test, the confirmatory tests were all performed with agitation. The factors studied were the SMBS and HAsc concentrations (High and Low) and pH (4 and 6). The filtrate was then analyzed using the XRF and the results are shown in Table 13.

EE —————————— reer : 27

He 1 o [ o [ oo [ 0 we oo | o [ o [ 0 [ww | me | Ww | wm

The readings in Table 13 were then analyzed by ANOVA Half Factorial, the result is shown in Table 14. The result confirmed the previous findings that HAsc concentration is a significant factor. The ANOVA test however does not state the relationship between the factor and the result. It only states the factors that are considered significant.

Table 14. Results of the half-factorial ANOVA - [Vode | | 3 | 247s | a6 | 463 | 002 | - [Grew [ 3 | was | Tee | 4@ | oom =

BB] 1 [ 7806 | 7806 | 453 | 0050 (Eo | | 2 | 7s | ee

Tota | ] 5 [| 44 | 00] 0] =

The control dosage of SMBS and HAsc at 0.00167M and 0.1M, respectively = where applied in pregnant solutions, and the composition of the precipitate and © s strip/barren solution after filtration are shown in Table 15. The major & composition of the precipitate was calcium at 96.5% and preceded by gold at - merely 1.952%.

Table 15. XRF analysis result of precipitation, ppm [G | feoass | Nb | fii0035 _

B25 | No | fe7 16 | MW | 14 (Pp [ 09 [| No [ NO

Further investigation on the dosage of SMBS and HAsc was performed with the objective of eliminating the high impurity of calcium in the precipitate. Table 16 shows the elemental composition in the precipitate at different dosage composition of SMBS and HAsc. An increase of dosage of 0.005M SMBS and 0.3M HAsc was able to eliminate the calcium impurity, and increased purity of gold which is important in the fusion with borax. Decreasing elemental species in the precipitate = is an advantage in gold fusion employing borax due to the decrease in the required i» heat energy to produce gold bead. = s Table 16. Precipitate composition at different SMBS and HAsc concentration = 0.005M 0.00167M 0.005M

SMBS, SMBS, SMBS, = 0.1M HAsc 0.3M HAsc 0.3M HAsc I” ~

A 1.5-liter volume of pregnant leach solution with Au content of 20ppm and initial pH of 5.6 was precipitated using SMBS and HAsc solutions. 1.2 liters of 0.005M

SMBS solution was added first, then after few minutes, 1.2 liters of 0.3M HAsc solution was added in the pregnant leach solution under stirred condition. It took a few minutes after the addition of HAsc solution for a visible formation of black precipitates. After two hours of stirring, the precipitation solution was filtered. The

XRF analysis of the precipitate and barren solution is shown in Table 17. The precipitate was 65.8% composed of gold with 1,199ppm concentration and the rest is composed of Fe, Sn, Pb and Sr. The filter paper containing the precipitates was oven dried to remove excess moisture. It was soaked with HCI solution (15%) and heated to boil until filter paper was degraded. The solution was filtered and washed with water, then dried. The XRF analysis of the precipitates and barren solution is shown in Table 17. The weight of the dried precipitate was 0.282 g with a recovery of 56%.

Table 17. XRF analysis of precipitant and barren solution, in ppm = [Ca [| ND | 145117 | i. -

The preferred embodiment of this invention is described in the above-mentioned . detailed description. It is understood that those skilled in the art may conceive = modifications and/or variations to the embodiment shown and described therein. on

Any such modifications or variations that fall within the purview of this description are intended to be included therein as well. Unless specifically noted, it is the intention of the inventors that the words and phrases in the specification and claims be given the ordinary and accustomed meanings to those of ordinary skill in the applicable art. The foregoing description of a preferred embodiment and best mode of the invention known to the applicant at the time of filing the application has been presented and is intended for the purposes of illustration and description. It is not intended to be exhaustive or to limit the present invention to the precise form disclosed, and many modifications and variations are possible in the light of the above teachings.

ENHANCED METHODS OF EXTRACTING PRECIOUS METALS -

AND METHODS OF TESTING pd

TECHNICAL FIELD OF THE INVENTION _

This invention pertains in general to the extraction of precious minerals and more oo particularly to the extraction of gold using an enhanced gravity concentration- - flotation, hypochlorite leaching and stage precipitation and methods of testing the - extraction processes. w

BACKGROUND OF THE INVENTION =

Small-scale miners of precious metals like gold in developing countries, because of lack of sufficient capitalization, often resort to the use of mercury (through amalgamation) to extract gold and reprocessed the mine tailings with the use of ~ cyanide (through cyanidation). This method exposes mercury and a possible transformation to methylmercury, an organic mercury which is more lethal.

Methylmercury is a neurotoxin that adversely affects the environment and health of living organisms.

Although cyanidation is an attractive procedure for small-scale miners due to its simplicity and its non-dependence to special equipment, the laborers are inevitably exposed to a chemical that destroys each and every cell it comes in contact to by cell-wall destruction. Studies have pegged that the maximum allowable cyanide exposure for humans is 0.2 ppm in water and 10 ppm in air.

The typical cyanide dosage used in cyanidation is around 2 ppm, which is more than enough to cause adverse effects to the body. Also, when the pH level of cyanide reaches near neutral (cyanidation is conducted at basic conditions), then I hydrogen cyanide is formed, which can cause poisoning by inhalation. -

Amalgamation, or the recovery of gold by allowing it to form an alloy with mercury called amalgam, is also a method-of-choice by small-scale mining operations. -

Using this process, gold is liberated enough to allow mercury to come in contact. 3

Gold can then be separated from mercury by applying heat to volatilize the low- i” vapor pressure mercury. The gold remaining is then smelted to yield the bouillon. -

Amalgamation uses about 0.5 Ibs mercury per metric ton of gold ore. However, - the health trade-off of the laborers is also apparent, as the final recovery of gold - from amalgam requires the mercury to be volatilized. Mercury vapor has long oi been deemed as a toxic substance that may cause poisoning should a person be exposed to vapors at 0.0002mg/m3, or ingest the mercury at 0.002 mg/kg body weight/day. The effects include weakness, respiratory illness, and at higher levels, even death.

US5232490A (490) discloses an “Oxidation/reduction process for recovery of precious metals from M\O; ores, sulfidic ores and carbonaceous materials.” discloses a process for separating precious metals from an MO, sulfidic or carbonaceous refractory ore or refractory feed such as tailings. The process of ‘490 includes the step of leaching a feed with a leach liquor that includes an acid selected from the group of HCI and H2SOs in the presence of MiO; and a reductant. A source of chloride ion is added to the leach sufficient to dissolve at least about 50% of the precious metals present in the ore. A portion of the leach is removed and precious metals are recovered from the removed portion. A portion of the chloride carrier is recycled to the leach to carry chloride values to the leach. In one embodiment, HCl is regenerated by pyrohydrolysis, which - minimizes harmful waste products. According to ‘490, its process can advantageously avoid the use of noxious reagents. D1 also teaches the . processes of gravity concentration, flotation, precipitation and treatment of = tailings. Among the objectives of ‘490 is to control the release of cyanide - compounds into the environment. However, the process of ‘490 is complicated I. and therefore costly as it targets primarily big mining companies as its users. =

This system provides a cheaper method of extracting precious metals from ores - without using any mercury or cyanide. This is done by enhanced gravity concentration-flotation, hypochlorite leaching and stage precipitation processes of extracting the metals from the ore. Moreover, wastewater from the extraction stage is treated before being discharged to the environment using a combination of zeolite and cocopeat, materials that are readily available to small-scale miners.

Methods of testing the foregoing are likewise provided.

BRIEF DESCRIPTION OF THE DRAWINGS

Figure 1 is a block diagram of the processes taught by the invention.

Figure 2 is a diagram of the crushing and grinding process.

Figure 3 is a diagram of the concentration process.

Figure 4 is a diagram of the dewatering process.

Figure 5 is a diagram of the extraction process.

Figure 6 is a diagram of the tailings and wastewater treatment process.

Figure 7 is a top view of the combination of the tailings settling pond and oO wastewater compartment. =

DETAILED DESCRIPTION OF THE INVENTION i. s Referring to Figures 1 to 7, the system is disclosed by this invention comprising of poe five processes namely crushing and grinding 1, concentration 2, dewatering 3, = extraction 4, and tailings and wastewater treatment 5. ”

In crushing and grinding process 1, the ore undergoes size reduction. The ore is ® received and goes to a jaw crusher 10 and the product through belt conveyor 11. ~

Any oversized ore (coarser than 1/2 inch) from a screen 12 goes to a roll crusher 13, while the undersized (finer than 1/2 inch) ore goes to a fine ore bin 14. The crushed ore with water are then fed into a ball mill 15 to achieve a product size with 80% passing 75 um. The product is then pumped to a hydrocyclone 16. The 15s underflow (coarser than 75 um) is recycled back to ball mill 15 while the overflow (75 um and finer) flows to a Falcon feed tank 17.

Falcon feed tank 17 feeds a Falcon gravity concentrator 21. Concentration in the

Falcon takes about 30 minutes per cycle (approximately 1 MT per cycle). The concentrate that contains the free gold is then collected in buckets. After which, the concentrate is fed to a table concentrator 22 for further cleaning. Table 22 produces a cleaner gold concentrate G, which is collected by a customized vacuum (not shown). The tailings GCT from Falcon gravity concentrator 21 and table 22 go to a flotation feed thickener 23. The pulp from flotation feed thickener 23 is pumped to a flotation feed conditioning tank 24. The pulp is mixed thoroughly inside conditioning tank 24 and the percentage of solids is maintained at 40%. The pH is adjusted to pH 9 with the addition of lime. The reagents used ro are CMS, a thionocarbamate collector, and Interfroth (IF) 6500, a glycol-based = frother. After a total of 30 minutes conditioning time, the pulp is fed to rougher- > cells 25a. The rougher concentrate RC flows to cleaner cells 26 while the rougher on tails go to scavenger cells 25b. The scavenger concentrate SC is also fed to ~ cleaner cells 26. The cleaner concentrate is pumped to the re-cleaner 27 for - further cleaning. The re-cleaner concentrate contains the gold associated and/or locked in sulfide minerals and becomes the final flotation concentrate FFC. The = cleaner tails and re-cleaner tails RCT is recycled back to rougher cells 25a. On = the other hand, the cleaner tails and scavenger tails become the final flotation - tails FFT and go to final tails thickener 28.

The extraction of metallurgy employs a oxidation and leaching processes.

Oxidation is necessary to convert the sulfide minerals to oxide before leaching. A mixture of solution containing sodium hypochloride, calcium hypochloride, sodium chloride and sodium hydroxide is first fed into the oxidation/leaching tank 42. The solution is mixed thoroughly to achieve homogeneity before adding final flotation concentrate FFC and shaking table concentrate. After 1.5 hours, calcium hypochloride is added. After 3 hours, pH is adjusted to pH 9 using hydrochloric acid. When the desired pH is achieved, calcium hypochloride is again added. The addition of calcium hypochloride is done every 10 minutes until the 4th hour. After the 4th hour, oxidation/leaching tank 42 is emptied. The mixture goes to filter press 32 for dewatering. The pregnant solution PS that contains the leached gold is collected in precipitation container 44 via a launder (not shown). Gas coming from the oxidation/leaching tank 42 is processed inside a gas scrubber 43 where - it is treated and clean air is released as a result. Sodium chloride is a by-product - of scrubber 43. - 5s The extraction process for gold employed is a 2-stage chlorination process. The i. process uses calcium hypochlorite as the hypochlorite-bearing reagent, sodium = chloride to stabilize the gold-chloride complex, and caustic soda and hydrochloric i. acid as pH-modifying reagents. The first stage is a 3 - 4 hours oxidation stage - wherein the sulfides minerals in the flotation concentrate is oxidized in alkaline = conditions. The pH is adjusted to 9.5 at the start and it is monitored throughout I. the oxidation process. Caustic soda is added to maintain the pH to 9.5. In this - stage, partial dissolution of exposed gold already occurs. Following the first stage is the second stage where the gold exposed from the oxidation stage is actively dissolved in neutral pH conditions. Here, the pH is adjusted and maintained around the range of 5-7 for 4 hours, with weighed amounts of calcium hypochlorite being added at 10-15 minute intervals. After the extraction process, the aqueous solution that contains the dissolved gold is separated from the solid residue and it will undergo the precipitation process to recover the gold.

To precipitate the gold, sodium metabisulfite is added in precipitation container 44. The solution is mixed thoroughly for at least 5 minutes. Thereafter, ascorbic acid is also added. The solution is again mixed thoroughly for at least 5 minutes.

The solution is left for about 2 hours to precipitate the gold. The solution is then filtered using pressure filter 45 to separate gold precipitates GP.

Claims

CLAIMS =

1. A method of extracting precious minerals comprising the steps of: =

A. crushing and grinding of ore; .

B. concentrating the crushed ore; -

C. dewatering of the concentrate; ~

D. extracting the precious metals from the concentrate; and +

E. treating the tailings and wastewater. =

2. The method according to Claim 1 wherein the following concentration - parameters for size of the feed (in terms of the percent passing Peso) and - the gravitational force of the concentrator (G Force) are: =o (Groce | 0 | 00 | 180

3. The method according to Claim 2 wherein 10 kilograms of ore was used as feed and was set at 10% solids.

4. The method according to Claim 2 wherein the average head assay was

7.05 gpt of gold.

5. The method according to Claim 2 wherein the operating water fluidization pressure for the concentrator is 6 psi.

6. The method according to Claims 2 to 5 wherein the optimal parameters of recovery are as follows: mm | iw | os | see | ds