WO2019203667A1

WO2019203667A1 - Enhanced methods of extracting precious metals and methods of testing

Info

Publication number: WO2019203667A1
Application number: PCT/PH2019/000005
Authority: WO
Inventors: Herman D. MENDOZA
Original assignee: University Of The Philippines Diliman
Priority date: 2018-04-18
Filing date: 2019-04-12
Publication date: 2019-10-24
Also published as: PH12018050174A1

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

ENHANCED METHODS OF EXTRACTING PRECIOUS METALS AND METHODS OF TESTING

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

BACKGROUND OF THE INVENTION

Small-scale miners of precious metals ike gold in developing countries, because of lack of sufficient capitalization, often resort to fire use of mercury (through amalgamation) to extract gold and reprocessed the mine tarings with the use of cyanide (through cyanidation). This method exposes mercury and a possible transformation to methytmercury, 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 smaS-scale miners dim to its simplicity and its non-dependence to special equipment the laborers are inevitably exposed to a chemical that destroys each and every ceB it comes in contact to by cefl-wall destruction. Studies have pegged that the maxknum 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 hydrogen cyanide is formed, which can cause poisoning by inhalation.

Amalgamation, or the recovery of gold by afiowing 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. Gold can then be separated from mercury by applying heat to volatflize the kw- vapor pressure mercury. The gold remaining is then smelted to yield the bouillon. Amalgamation uses about 0.5 lbs 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

toxic substance that mrrvciMi_Sftiwisoninn shoutri a person he exposed to vapors at G.0002mg/ 3, 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 MnO_å ores, suffktic ores and carbonaceous materials.' discloses a process for separating precious metals from an MhO_¾ sutfidic 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 add selected from the group of HCI and tfeSG* in the presence of

and a reductant A source of chloride ion is added to the teach sufficient to (fissohre 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 cany chloride values to

the leach. In one embodiment, HCI is regenerated by pyrohydrotysis, which minimizes harmful waste products. According to ^*490, its process can advantageously avoid the use of noxious reagents. D1 also teaches the processes of gravfty concentration, flotation, precipitation and treatment of taffings. Among the objectives of ‘490 is to control the release of cyanide compounds into the environment However, the process of‘490 is oompficated and therefore costly as it targets pranarily big mining companies as its users.

SUMMARY OF THE INVENTION

This system provides a cheaper method of extracting precious metals horn 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 toe ore. Moreover, wastewater from toe extraction stage is treated before being rflscharged to toe environment using a combination of zeofite and cocopeat, materials that are readily avaflable 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 toe concentration process.

Figure 4 is a diagram of toe dewatering process.

Figure 5 is a diagram of the extraction process.

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

Figure 7 is a top view of the combination of the taiings setling pond and wastewater compartment.

DETAILED DESCRIPTION OF THE INVENTION

Referring to Figures 1 to 7, the system is disclosed by this invention comprising of five processes namely crushing and grindmg 1, concentration 2, dewatering 3, extraction 4, and taings 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 be* conveyor 11.

Any oversized ore (coarser than 1/2 inch) from a screen 12 goes to a roB crusher 13, «ride 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 pen. The product is then pumped to a hydrocydone 16. The underflow (coarser than 75 pm) is recycled back to ball mffl 15 while the overflow (75 pm and finer) tows to a Falcon feed tank 17.

Falcon feed tank 17 feeds a Falcon gravity concentrator 21. Concentration in fire Falcon takes about 30 minutes per cycle (approximately 1 MT per cycle). The concentrate that contains the free gold is then coiected 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 taffings 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 nixed

thoroughly inside conditioning tank 24 and the percentage erf sofids is maintained at 40%. The pH is adjusted to pH 9 with the addition of Gme. The reagents used are CMS, a thionocarbamate collector, and Interfroth (IF) 6500, a glycol-based frother. After a toted of 30 minutes conditioning time, the pulp is fed to rougher- cefls 25a. The rougher concentrate RC flows to dealer cefe 26 while the rougher tads go to scavenger ceDs 25b. The scavenger concentrate SC is also fed to cleaner cefe 26. The deaner concentrate is pumped to the re-deaner 27 for further cleaning. The re-deaner concentrate contains the gold associated and/or locked in sulfide minerals and becomes the final flotation concentrate FFC. The deaner tails and re-deaner tails RCT is recycled back to rougher cells 25a. On the other hand, the deaner tads and scavenger tads become the final flotation tails FFT and go to find tads thickener 28.

The extraction of metallurgy employs a oxidation and leaching processes, Oxidation is necessary to convert the suffide 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 a d. When the desired pH is achieved, calcium hypochtorkte is again aided. The addition of caldum hypochloride is done every 10 minutes until the 4th Iwxir. 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 b processed inside a gas scrubber 43 where it b treated and clean sir b released as a result Sodium chloride b a by-product of scrubber 43. The extraction process for gold employed b a 2-stage chlorination process. The process uses calcium hypochlorite as the hypochlorite-bearing reagent, sodium chloride to stabilize the gold-chloride complex, and caustic soda and hydrochloric acid as pH-mocSfying reagents. The first stage b a 3 - 4 hours oxidation stage wherein the sulfides minerals in the flotation concentrate b oxidized in alkaline conditions. The pH b adjusted to 9.5 at the start and it b monitored throughout the oxidation process. Caustic soda b added to maintain the pH to 9.5. In this stage, partial dissolution of exposed gold already occurs. Foflowmg the first stage b the second stage where the gold exposed from the oxidation stage b actively dissolved in neutral pH conditions. Here, the pH b adjusted and maintained around the range of 5-7 for 4 hours, with weighed amounts of calcium hypochlorite being added at 10-15 minute intervab. After the extraction process, the aqueous solution that contains the dissolved gold b separated from the sofid residue and it will undergo the precipitation process to recover the gold. To precipitate the gold, sodium metabbu!fite b added in precipitation container 44. The solution is mixed thoroughly for at least 5 minutes. Thereafter, ascorbic add b also added. The solution b again mixed thoroughly for at least 5 minutes. The solution b left for about 2 hours to pretipftate the gold. The solution b then filtered using pressure filter 45 to separate gold precipitates GP.

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

The tailings horn find tails thickener 28 is discharged into final tailings setling 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 taithgs settling pond 51 and the barren sofids BS from filter press 32 and pressure filter 45 can be packed into sacks 56 or simaar containers for final disposal or may be utilized as addffional support structures to counter erosion or even for landscaping.

The washings aid 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 botom of cocopeat-zeoSte layer 54 goes to a holding area/empty compartment 55. When this compartment becomes full, it overflows to second compartment 532 with a cooopeat-zeofite layer for the second stage of treatment. The treated water from the second stage proceeds to third compartment 533 with 100% zeolite for the fmal stage of treatment The treated water from the third stage, together with dear water CW, is then discharged to river R.

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The recovery of free gdd was conducted through the gravity concentration provided by this invention. Gravity concentration tests were done using an L40 Falcon gravity concentrator. A 324actorial experiment design was performed to determine the operating parameters, particularly the size of the feed n terms of the percent passing Peo) and the gravitational trace of the concentrator (G Force). Both parameters were set at three levels as shown in Table 1.

Table 1. Design of experiment for the Fakon gravity concentrator

For each ran, 10 kSograms of we 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 fluidzation pressure for the concentrator as also maintained at 6 psi. The resuts are presented in Table 2.

Table 2. Results of the gravity concentration experiments.

The results in Table 2 show variable yield for the different setings of Pm 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 erf recowry versus G Force is illustrated in Figure 2. A general trend can be observed wherein finer partide sizes require lower G Forces to attain higher recoveries.

Figure 2. Recovery versus G Force.

Flotation Tests

A Taguchi design of experiment (DOE) was performed to deternroie the parameters for flotation. The parameters considered were feed size (PM), 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 TagucN design of experiment set for notation tests

For the colector and frother machines used. CMS is an Australian brand of ihionocarhamate collector. SIBX is a ioca!y produced xanthate and S701 is a Nasaco dithiopbosphate colector. Nasfiroto 240 is a glycol based frother, Nasfroth 626 is an alcohol based frother both coming from the Nasaco braid, 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 «ms conducted in a 5L flotation cel at 40% sofids. 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 Iran toe froth phase into a sample pan for 1.5 minutes. The aeration is then tinned off, another dose of toe reagent is added aid a new pan is prepared to the next stage before turning the aeration on again. R2 to R4 followed a scraping time of 2 minutes, also following afl 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. The experiment highfighted the significant factors foal affected the gold flotation conoentrate grade and recovery of the ore sample namely, collector dosage and pH. A summary of the flotation runs that yielded significant results is given in Table 4.

Tabfe 4. Flotation run resu s

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 to 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 afl particle sizes considered yielded similar concentrate

grade and recovery. A beter visualization of the relationship between the different factors is illustrated in Figure 3.

Figure 3. Relationships of the differert factors.

Flotation time was determined to be between the 5.5 to 7.5 minutes period. A 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 pflot-sca!e.

Figure 4. Recoway curve from the batch flotation test.

Extractive Metallurgy Tests

The recovery of gold from flotation concentrate and purification of gold from the gravity concentrator were conducted through leaching and precipitation. An extraction process was developed that fu!ly deviates from the use of cyanide and mercury. Several laboratory tests have been conducted to estabfish the extraction procedure and operating parameters. The laboratory tests resulted in about 80% recovery of gold. a. Uxiviant Composition

Preliminary experiments were conducted to determine tee type erf lixiviant to be used for leaching. The types of fix ants compared were thiourea, thiosulfate, iodine, and chlorine (see Figure 5.). The results showed that iodine yielded tee 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 gdd leaching (via chlorination process).

is

Fyum 5. PfeBnmayexperimeifccompamg various types of Mviants.

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

7a6fe5. Types and dosages ofBxiviants for leaching

The highest recovery, about 60%, was achieved using redpe B. This is due to the differing properties of the completing ligands such that the calcium hypochlorite is more stable than sodium hypochlorite and has a higher chlorine content white sodium hypochlorite has kinetics twee that of calcium hypochlorite. Based on this, the succeeding leaching experiments used the following recipe: 18% tfcO, 8% NaCI, 58% NaOCI, 16% HCI, and 3% CafOCfc.

A further test involving decreased amount of hydrochloric arid was conducted, it was found that decreasing the hydrochloric content up to 0.32% of the original concentration cam 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 cfesotution of gold in chlorination is oxidation of the gold, the sulfide minerals would abo 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 add needed for hypochlorite oxidation. The decrease in the amount of hydrochloric acid was based on foe resulting pH of the water-ptus-hydrochioric add mixture at the start of the leaching process. Figure 7 shows the result of HC1 dHuSon on toe recovery of gold

Figure 7. Effect of HCldlut n on the recovery of gold. s b. Mode of Addition of Lixiviant

Gold dissolution occurs in two stages. (Au+J is an intermediate species in the leaching reaction. The (Au+J that was irtiialy formed is the species would react wti the Cl- ions to form the more stable gold chloride [AuCI4i complex. However, if the concentration of |Air] increases rapidly, based on Le^* ChateEer’s principle, the 10 backward reaction would take place. This wotdd slow down the rate of dissolution of gold and would lessen the recovery of the process. To ensure the continuous

oxidation of gold from the solids, the [Air] concentration should be kept at a minimim This can be done by ensuring that the fAu+1 formed would immediately react with Cl- ions in the solution. The interrrattert adtftion of Ca(OCI)2 and NaOCI ensures that a steady supply of O- are avaiaUe to react wih the Air ions that are formed throughout the teaching process. Bulk addition of the hypochlorite spedes might cause the amount of Ct- ions avaSable to drop during the latter stages of the process due to compfescation of other metals to the chlorine ions. The design of experiment and the results are summarized in Table 6 and Figure 8. Bulk adcflion is adding afl the leaching reagents at toe start and wanting for a fixed amount of time for the reaction to finish. The K oG method is an incremental mode of adcfition.

It tovdves adding HCI at the start, and toen NaOCI is added incrementally untfl the end of leaching time.

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

c. Pulp Density

Tests were done to determine the pulp density that should be used in leaching. The result Musirated in Figure 9 show that higher leaching recoveries were observed at lower puip densities. Thfe could be Aie to beter kinetics since ft 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 effecthrity of the precipitation process as shown in Figure 7. To achieve an optimum leading recovery and its subsequent precipitation, the pulp density should be at 1.1421 g/mL

d. Roasting as Pirtreatment and Die 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 stdfide minerals. Based on tests conducted (results are shown HI Table 7), roasting was found to be a significant factor to foctease gold recovery. The tests also served as verification on the significance of sodum chloride in the process. It was found that sodium chloride does not significantly affect the recovery thus, can be omitted to decrease reagent cost

Table 7. Statistical result of the ANOVA futi factorial analysis

From Table 8, it was observed that the mixture of IbdvianJ contributed to the increase in recovery. The effect of using cfifferent forms of hypochlorite, particularly sodium hypochlorite (NaOCI) and calcium hypochlorite (CaOCh) 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 m¾or considerations in determining the effective pulp density for leaching whte keeping dilution at the minimum. Based on foe results, the NaOCI + Ca(OCl)2 mix resulted in foe highest recovery. Therefore, foe fcwiant mix used in foe succeeding leaching tests comprised of 48% NaOCI and 2% Ca{OCI)2 of the fixiviant volume.

7aWe 8. Effect of pnetreatment, hypochlorite form and presence of sodium chloride

e. Leaching Time

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 cfissolution of gold occurred alter 3 - 4 hows of leaching. This is because gold is ooddized before the competing sufphkSc minerals. However, prolonging leaching beyond 4 hours decreased foe 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 minerals. Based on the observed trend, leaching time was preferred to be at 4 hours.

Figure 10. Effect of leaching time on the recovery of gold.

f. Simultaneous Oxidation and Leaching

Due to S02 emissions from roasting, an alternative pretreatment process was examined. Several oxidizing agents were considered. Add washing using hydrochloric add (HCI) was explored. However, as determined through Bruce method, no gold is locked within the carbonates of the ore. HCI is only effective in oxidizing carbonates and has no effect on suiphidic minerals. Sodium hydroxide (NaOH) or caustic soda was then shxfied. Previous stixfies 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

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 Igoii method. It involves a two-part process: 1) oxidation of sulfides at alkaline conditions to expose the locked gold (as weB as leach any free gold in the concentrate); and 2) leaching of the exposed gold. In this method, an amount of NaOCt, NaCl, NaOH, and Ca(OCI)2 is first mixed until homogeneity, The Ca(OCI)2 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(OCI)2, 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 120°' minute (2^nd hour) but decreases as leaching time is further increased. This may be due to the gold complex becoming unstable after the 120*' minute. The

reagent NaCI increases the stability of the complex and the amount of NaCI at the 120^ft 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 T able 10.

Figure 11. Recovery of gold using the simuiatneousoxidaiion and leaching method at various leaching times. Table 10. Modified Operating Parameters for Pre4reatment and Leaching

Note that NaCI and Ca(OCf)2 are added incrementaily to maintain the strength of HOC1 spedes in the solution as well as the stability of the gold auro co pter formed. Incremental addition was done every ten minutes until the 4^m hour. g. Precipitation Stage precipitation was performed using sodium metabisulfite (SMB S) and ascorbic add (HAsc). Initial tests were already conducted to determine the feasibility of this process for gold recovery. These tests were successM in precipitating go!d from pregnant solution after leaching. Using a blowtorch, precipitates obtained were converted into gold beads thereby verifying the viability of the process. W¾h 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

A tower reading of the XRF is favored because it impfies more gold was precipitated and only a small amount remaned 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 add (HAsc) since within this pH range, it is in its deprotonated form. Hie deprotonated form of HAsc is a stronger reductant as compared to its other species. The data in Table 11 were tested for analysis of variance (ANOVA) at full factorial to determine the parameters which has significant effect on gokf recovery. The result is shown in Tabfe 12, where factor A is the SMBS concentration, B is HAsc 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 foe 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.

Table 12. ANOVA table showing the factors and their interactions that affect the result of precipitation.

Table 13. Gold concentration of the stripped solution at ₍fflerent conditions.

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 HAse concentration is a significant factor. The ANOVA test however does not state the relationship between the factor and the result It only states flie factors that are considered significant

Table 14. Results of the hatMactorialANOVA

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 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

Further investigation on the dosage of SMBS and HAsc was performed wBh foe ol^ecfive of eliminating the high impurity of calcium in the precipfete. Table 16 shows the elemental composition in foe precipitate at cKfferent 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 heat energy to produce gofd bead.

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 0.0Q5M 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 boit until filter paper was degraded. The solution was filtered aid 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%.

The preferred embodiment of this invention is descrfeed in the above-mentioned detailed description. It is understood that those skiBed in the art may conceive mocfifications and/or variations to the embodiment shown and described therein. Any such modifications or variations that taO within the purview of this description are intended to be included therein as weR. Unless specifically noted, it is toe intention of toe inventors that the words and phrases in toe specification and claims be given the ordinary and accustomed meanings to those of ordinary skill in the applicable art The foregoing description of a pretoned embodiment and best mode of toe invention known to the applicant at toe time of fiing 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 toe precise form disclosed, and many modifications and variations are possible in the light of toe above teachings.

Claims

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

A. crashing 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 Clam 1 wherein the following concentration parameters for size of the feed (in terms of the percent passing PM) and the gravitational force of the concentrator (G Force) are:

3. The method accordmg to Clam 2 wherein 10 kiograms 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 pst.

6. The method according to Claims 2 to 5 wherein the optimal parameters of recovery are as follows: